THE  METALLURGY  OF 
THE  COMMON  METALS 

Gold,  Silver,  Iron  (and  Steel), 
Copper,  Lead  and  Zinc 


BY 
LEONARD  S.  AUSTIN 

Formerly  Professor  of  Metallurgy  and  Ore  Dressing, 
Michigan  College  of  Mines 


FIFTH    EDITION 
REVISED  AND  ENLARGED 


NEW  YORK 

JOHN  WILEY   &  SONS,   INC. 

LONDON:  CHAPMAN  &  HALL,  LIMITED 

1921 


•rr 


COPYRIGHT 

First     Edition,  MINING  &  SCIENTIFIC  PRESS,  1907 
Second        "          :    "  "  "       1909 

Third  .*  1  tf  t  '    :   "  "  "       1911 

Fourth     '  "  '      "A  "  "       1913 

Fiftn/  ',t  ;  "  ;  .  LE.OFARD  S    AUSTIN  1921 


BRAUNWOPTH    IL   CO. 

BCXJK    MANUFACTURERS 

BROOKLYN.   N.   Y. 


PREFACE   TO  THE  FIFTH  EDITION 


SINCE  1913,  the  date  of  the  last  edition,  such  radical  changes  and 
improvements  have  been  made  in  the  metallurgy  of  the  common  metals, 
that  this  edition  of  1921  has  been  largely  rewritten  to  bring  it  in  accord  with 
present  practice,  as  will  be  seen  by  examination  of  the  following  pages. 
Great  pains  have  been  taken  to  clearly  set  forth  underlying  principles  and 
at  the  same  time  to  give  the  details  of  methods  and  of  metallurgical  equip- 
ment, and  their  cost.  It  is  realized  that,  due  to  the  rapid  advance  in  prices, 
the  costs  of  operation  have  lately  been  subject  to  serious  modification.  A 
chapter  has  been  devoted  to  questions  of  the  economic  situation  of  the 
business  of  metallurgy.  Little  attempt  has  been  made  to  describe  methods 
not  now  in  use. 

L.  S.  AUSTIN. 
Los  ANGELES,  May  1,  1921. 

PREFACE  TO  THE  FIRST  EDITION 


THIS  outline  of  the  metallurgy  of  the  common  metals,  namely,  gold, 
silver,  iron,  copper,  lead,  and  zinc,  is  devoted  to  the  description  of  processes 
for  winning  these  metals  from  their  ores  and  then  refining  them.  The 
metallurgy  of  iron  is  treated  only  to  the  point  where  pig-iron  is  obtained. 

Following  the  description  of  ores,  as  well  as  of  the  fuels  used  in  smelting 
them,  and  the  materials  of  which  the  furnaces  are  constructed,  we  come  to 
the  sampling,  for  the  determination  of  the  exact  value  of  the  ore  before 
treatment. 

A  chapter  has  been  devoted  to  the  subject  of  thermo-chemistry  as 
applied  to  igneous  methods  of  extraction.  The  winning  or  reduction  of  the 
various  metals  is  then  taken  up  in  order,  and  is  followed  by  a  description 
of  the  methods  of  refining  them.  Attention  is  then  given  to  commercial 
considerations,  since  the  processes  must  be  conducted  in  a  profitable  way. 

The  author  is  indebted  to  Mr.  F.  L.  Bosqui,  who  has  not  only  read 
the  manuscript,  but  has  modified  the  portion  devoted  to  the  cyaniding 
of  gold  and  silver  ores,  as  his  special  knowledge  has  justified.  For  the 
subject  matter  relating  to  the  smelting  of  silver-lead  and  copper  ores,  the 

iii 

447637 


iv  PREFACE 

author  has  drawn  on  his  own  experience,  gained  during  a  quarter  of  a 
century  of  practical  work. 

L.  S.  AUSTIN. 
HOUGHTON,  May  1,  1907. 

PREFACE  TO  THE   SECOND   EDITION 


THE  experience  gained  in  using  the  first  edition  has  suggested  many 
changes,  and  the  book  has  accordingly  been  re-written,  adding  new  matter, 
describing  other  processes,  and  keeping  step  with  modern  practice. 

In  Part  I  the  subject  of  thermo-chemistry  has  been  expanded,  and  a 
table  of  heats  of  formation  given.  The  description  of  the  cyanide  process 
has  been  amplified  and  brought  up  to  date,  for  milling  methods  are  being 
rapidly  improved,  and  cyanidation  is  having  increased  application,  espe- 
cially in  the  treatment  of  silver-bearing  ores.  The  metallurgy  of  zinc  has 
been  treated  more  fully,  and  particular  attention  given  to  the  principles 
underlying  the  smelting  of  zinc  ores. 

In  the  part  devoted  to  refining  there  has  been  added  the  making  of 
wf ought-iron  and  steel,  the  refining  of  zinc,  and  the  electrolytic  refining  of 
lead. 

Plant  and  equipment  is  placed  in  a  separate  chapter,  while  the  division 
describing  the  economics  of  metallurgy  has  been  thrown  into  a  more 
systematic  form.  The  author  is  indebted  to  Mr.  E.  A.  Hersam,  who  read 
the  manuscript  of  the  second  edition  and  made  numerous  suggestions  and 
corrections. 

The  author  is  indebted  to  the  following  companies  for  the  use  of  certain 
of  the  illustrations  in  this  book:  Allis-Chalmers  Co.,  Milwaukee,  Wis.; 
Power  &  Mining  Machinery  Co.,  Cudahy,  Wis.;  Chisholm,  Matthew  & 
Co.,  Colorado  Springs,  Colo.;  F.  M.  Davis  Iron  Works  Co.,  Denver,  Colo.; 
Steams-Roger  Mfg.  Co.,  Denver,  Colo.;  Pacific  Tank  Co.,  San  Francisco, 
Cal.;  Redwood  Manufacturers  Co.,  San  Francisco,  Cal.;  Galigher  Ma- 
chinery Co.,  Salt  Lake  City,  Utah;  Traylor  Engineering  Co.,  Allentown, 
Pa.;  Blaisdell  Co.,  Los  Angeles,  Cal.;  Denver  Engineering  Works  Co., 
Denver,  Colo.;  Trent  Engineering  &  Machinery  Co.,  Salt  Lake  City,  Utah; 
Risdon  Iron  Works,  San  Francisco,  Cal.;  Colorado  Iron  Works  Co., 
Denver,  Colo. ;  Cyanide  Plant  Supply  Co.,  Ltd.,  London;  The  Jeffrey  Mfg. 
Co.,  Columbus,  Ohio. 

L.  S.  AUSTIN. 
HOUGHTON,  August  1,  1909. 


PREFACE 


PREFACE  TO   THE   THIRD  EDITION 


THE  present  edition  has  been  more  systematically  arranged,  and  errors 
have  been  eliminated.  Important  and  recent  changes  in  smelting  practice 
and  in  the  cyanidation  of  gold  and  silver  ores  have  justified  the  insertion  of 
additional  matter,  much  of  which  has  come  under  the  direct  observation 
and  inquiries  of  the  author. 

L.  S.  AUSTIN. 
SALT  LAKE  CITY,  UTAH,  March  1,  1911. 

PREFACE   TO  THE  FOURTH  EDITION 


IN  this  edition,  that  part  of  Chapters  X  and  XIV  which  discusses  the 
cyaniding  of  gold  and  silver  ores  respectively,  has  been  written  by  M.  W. 
von  Bernewitz,  formerly  of  the  Associated  Northern  and  Associated  Mines, 
Kalgoorlie,  Western  Australia,  and  now  on  the  staff  of  the  Mining  and 
Scientific  Press.  The  chapter  on  the  metallurgy  of  zinc  has  been  re-written 
by  Mr.  R.  G.  Hall,  long  manager  for  the  United  Zinc  &  Chemical  Co.,  and 
later  in  general  consulting  practice.  These  gentlemen  are  specially  quali- 
fied for  the  subjects  they  have  undertaken  and  have  incorporated  the 
recent  practice  in  the  art. 

L.  S.  AUSTIN. 
SALT  LAKE  CITY,  UTAH,  August  15,  1913. 


J 

^ 


*  •" 


(> 


CONTENTS 


PART  I— GENERAL  METALLURGY 

CHAPTER  I 
ORES  AND  METALS: 

PAGE 

Definition  and  Classification  of  Ores 3 

Methods  of  Treatment 4 

Classification  of  Metallurgical  Operations 5  * 

Principles  Relating  to  the  Refining  of  Metals 5 

Physical  Constants  (Meta  s  and  Gases) 6 

Moulding  and  Casting  Metals 7 

CHAPTER  II 

FUELS: 

Fuels : 11 

The  Natural  Solid  Fuels 12 

Producer  Gas 23 

Pulverized  Coal 29 

CHAPTER  III 
REFRACTORIES: 

Refractory  Materials  and  Their  Properties 31 

Acid  Refractories 32 

Neutral  and  Basic  Refractories 37 

CHAPTER  IV 
THE  PREPARAT.ON  OF  ORES: 

Principles  of  Sampling 39 

Receiving,  sampling,  crushing,  bedding  and  storing  ores 40 

Sampling  Metals 48 

CHAPTER  V 

CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING: 

Principles  of  Crushing 50 

Coarse  or  Primary  Crushing 51 

Intermediate  or  Fine  Crushing  or  Coarse  Grinding 55 

Fine  Grinding 64 

Screening 66 

Classifying 68 

vii 


viii  CONTENTS 

CHAPTER  VI 
METALLURGICAL  FURNACES: 

PAGE 

Shaft  Furnace 72 

Reverberatory  Furnace 73 

CHAPTER  VII 
COMBUSTION: 

Principles  of  Combustion 75 

Temperatures  of  Combustion 79 

CHAPTER  VIII 
METALLURGICAL  THERMO-CHEMISTRY  : 

Methods  of  Determining  Thermic  Values 84 

Heats  of  Formation  of  Chemical  Elements 86 

CHAPTER  IX 
ROASTING: 

Kinds  of  Roasting 88 

Chemistry  of  Roasting 89 

Roasting  Ores  in  Lump  Form;  Heap-roasting 92 

Roasting  of  Ores  in  Pulverized  Condition 94 

The  Long-hearth  Reverberatory  Roaster 95 

Mechanically-operated  Roasting  Furnaces 97 

Roasting  of  Matte .' 108 

Losses  in  Roasting 108 

Capacity  of  Furnaces  and  Cost  of  Roasting 109 

Blast  or  Pot-roasting 110 

Sinter  Roasting 110 

Triple  Roasting 113 

CHAPTER  X 

CONCENTRATION  OF  ORES  AS  A  SUBSIDIARY  OPERATION  IN  METALLURGY: 

Concentration 114 

Flotation , 116 

PART  II— GOLD 

CHAPTER  XI 
GOLD  ORES  AND  CLASSIFICATION  FOR  MILLING: 

Occurrence 121 

Valuation  of  Gold  and  Silver 122 

Classification  for  Milling 122 

CHAPTER  XII 

AMALGAMATION  : 

Stamp  Milling  with  Plate  Amalgamation 124 

General  Arrangement  of  a  Gold  Stamp-mill 130 

Concentration  in  Stamp-milling 131 


CONTENTS  ix 

CHAPTER  XIII 

HrDROMETALLURGY   OF   GOLD  ORES: 

PAGE 

Milling  Ores  in  Aqueous  Solution 133 

CHAPTER  XIV 
CHLORINATION  OF  GOLD  ORES: 

Ores  Suited  to  Chlorination 135 

The  Goldfield  Chlorine  Mill  Co.,  Goldfield,  Nev 136 

Barrel  Chlorination 137 

CHAPTER  XV 
CTANIDINQ  OF  GOLD  ORES: 

Outline  of  the  Process  of  Cyaniding 143 

Ores  for  Cyanidation .' 145 

Chemistry  of  the  Cyanide  Process  for  Gold  Ores 146 

The  Standard  Systems  of  Cyaniding 150 

Sand  Leaching • 151 

Double  Treatment 153 

Filter-Slime  Treatment 157 

Slime-Agitation 157 

Pneumatic  Agitators 158 

Mechanical  Agitators 159 

Combined  Mechanical  and  Pneumatic  Agitators 160 

Agitation  Treatment 161 

Continuous  Counter-current  Decantation 165 

Filtration  or  Separation  of  Metal-bearing  Solution  from  Slime 166 

Vacuum  Filtration 168 

Pressure  Filtration 171 

General  Remarks  on  Filters 175 

The  Crowe  Vacuum  Process 177 

The  Precipitation  of  Gold  from  Cyanide  Solutions 177 

The  Merrill  Precipitation  Process 178 

The  Zinc  or  Extractor  Box 181 

The  Clean-up 182 

Dryuag  and  Refining  the  Gold  Precipitate 184 

Refining  with  Bichromate 184 

Capital  Costs  of  Slime  Plants 186 

Costs  of  Dissolution  by  Slime  Agitation 186 

CHAPTER    XVI 
TYPICAL  GOLD-MILL  PRACTICE: 

Cyaniding  Free-milling  Porous  Ores : 189 

The  Wasp  No.  2  Mill,  South  Dakota 189 

Ores  of  Clayey  Nature  by  Cyaniding 189 

The  Victorious  Mill,  Western  Australia 190 

The  Kolar  Field 191 

The  City-deep  Mill 191 

The  Mill  of  the  Consolidated  Langlaaghte  Co.,  Rand,  South  Africa 193 

The  Homestake  Mill,  Lead,  South  Dakota 195 

The  Liberty  Bell  Mill,  Telluride,  Colo 199 


CONTENTS 

PAGE 

Treatment  of  Telluride  Ores 199 

The  Golden  Cycle  Mill 201 

The  Victor  Plant  of  the  Portland  Gold  Mining  Co 202 

The  Kalgoorlie  District,  Western  Australia 203 

The  Oroya-Brownhill  Mill,  Kalgoorlie  District 204 

The  Hollinger  Mill,  Porcupine  District,  Ontario 205 

The  Tom  Reed  Mill,  Oatman,  Ariz 207 

The  United  Eastern  Mill,  Oatman,  Ariz 209 


CHAPTER  XVII 

TREATMENT  OF  GOLD-MILL  CONCENTRATES: 

Classification 213 

The  Alaska-Treadwell  Concentrate  Treatment  Plant 215 

Concentrate  Treatment  at  the  Goldfield  Consolidated,  Goldfield,  Nev 218 

CHAPTER  XVIII 

VARIOUS  TREATMENTS  AND  CALCULATIONS: 

Flotation  and  Cyaniding 223 

Drying  and  Cyaniding 223 

Treatment  of  Tailings  from  Acid  or  Ammonia  Leaching 223 

Calculation  of  Tonnages  in  Mills 223 

CHAPTER  XIX 
SMELTING  GOLD  ORES: 

Blast-furnace-smelting  vs.  Cyaniding  of  Gold  Ores 226 

The  Price  of  Gold  Ores.     Also  the  Cost  of   Producing  and   Selling.     The 
Price  of  Gold 227 

PART  III— SILVER 
CHAPTER  XX 

SILVER,  ITS  ORES  AND  THEIR  TREATMENT: 

Characteristics  of  Silver  Ores 231 

Extraction  of  Silver  from  Ores 232 

Treatment  of  Silver  Ores . .  232 

CHAPTER  XXI 
AMALGAMATION  OF  SILVER  ORES: 

Wet  Silver  Milling  with  Tank-settling 235 

The  Boss  Process  of  Silver  Milling 243 

The  High-grade  Nipissing  Mill,  Cobalt,  Ontario 243 

Plate  and  Pan-Amalgamation  and  Concentration  of  Silver  Ores 244 

The  Chloridizing  Roasting  of  Silver  Ores 246 

Dry  Silver  Milling  (Reese  River  Process) 249 

The  Patio  Process..  .  249 


CONTENTS  xi 

CHAPTER  XXII 
SILVER  MILLING  BY  HYDROMETALLURGICAL  PROCESSES: 

PAOB 

The  Principles  of  the  Hydrometallurgy  of  Silver 250 

The  Augustin  Process 250 

The  Ziervogel  Process ' 251 

The  Hyposulphite  Lixiviation  Process  for  Silver  Ores 254 

The  Russell  Process 254 

CHAPTER  XXIII 
CYANIDATION  OF  SILVER  ORES: 

Principles  of  Cyanidation 256 

The  Precipitation  of  Silver  from  Cyanide  Solution 257 

The  Santa  Gertrudis  Precipitating  and  Refining  Plant 259 

Precipitation  by  Aluminum  Dust 261 

Drying  and  Refining  Silver  Precipitate 261 

Chemistry  of  the  Process  for  Silver  Ores 264 

Typical  Silver  Mills 265 

Belmont  Milling  Co.'s  Mill,  Tonopah,  Nev 265 

Cyanidation  of  Mixed  Silver  Ores  at  the  San  Francisco  Mill,  Pachuca,  Mex.  268 

The  Waihi  Grand  Junction  Mill,  Waihi,  N.  Z 270 

Milling  Practice  at  Cobalt,  Ontario 274 

The  Nippissing  Co.'s  "Low-grade"  Mill,  Cabalt,  Ontario 274 

CHAPTER  XXIV 

PARTING  GOLD-SILVER  BULLION,  PRICES  AND  COSTS: 

Parting  Gold-silver  Ingots  or  Bars  with  Acids 278 

Electrolytic  Parting  of  Gold  from  Silver 279 

Prices  and  Costs 280 

PART  IV— IRON  AND  STEEL 

CHAPTER  XXV 

IRON  ORES  AND  THEIR  SMELTING: 

Classification  and  Occurrence  of  Iron  Ores 283 

Roasting  Iron  Ores 285 

The  Agglomeration  of  Fine  Ores 286 

Smelting  for  Pig  Iron 287 

Iron  Blast-furnace  and  Plant 287 

Gas  Cleaning 287 

Hot-blast  Stoves 295 

Blast-furnaces  and  Accessories 298 

Operation  of  the  Blast-furnace 300 

Irregularities  of  Furnace  Operation 302 

Disposal  of  Slag  or  Cinder 303 

Disposal  of  Pig  Iron 304 

Dry-air  Blast 305 

Chemical  Reactions  of  the  Blast-furnace 306 

The  Heat  Balance  of  the  Blast-furnace 309 


Xll  CONTENTS 

PAGE 

Burdening  the  Blast-furnace 310 

General  Arrangement  of  the  Blast-furnace  Plant 312 

Pig  Iron 314 

Classification  of  Pig  Iron 314 

Influence  of  the  Contained  Elements  on  the  Character  of  the  Pig  Iron 316 

CHAPTER  XXVI 
WROUGHT  IRON  AND  STEEL: 

The  Manufacture  of  Wrought  Iron  by  the  Puddling  Process 318 

Steel  Making 320 

Steel  Making  by  the  Acid  Bessemer  Process 321 

The  Basic  Bessemer  Process 326 

Steel  Making  in  the  Open-hearth  Furnace 326 

The  Open-hearth  Reverberatory  Furnace ^ 327 

The  Acid  Open-hearth  Process '. .  .  '. 333 

Basic  Open-hearth  Process 334 

Calculation  of  Charge 335 

The  Open-hearth  Building 339 

The  Duplex  Process  of  Steel  Making 340 

Duplex  and  Electric  Furnace  Plant 341 

Electric  Steel-making 343 

Varieties  of  Steel 345 

Iron  Ore  and  Pig  Iron  Prices 347 

PART  V— COPPER 
CHAPTER  XXVII 

COPPER  ORES  AND  THEIR  TREATMENT: 

Characteristics  of  Copper  Ores 351 

Extraction  of  Copper  from  Its  Ores 353 

CHAPTER  XXVIII 

COPPER  BLAST-FURNACE  SMELTING  OP  OXIDIZED  ORES: 

Blast-furnace  Smelting  of  Oxidized  Ores 355 

Lake  Superior  Copper  Country  Smelting  of  Copper  Reverberatory  Slag 357 

Smelting  to  Black  Copper  by  the  Union  Miniere  du  Haut  Katanga 359 

CHAPTER  XXIX 

BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES: 

Matte  Smelting 360 

The  Messiter  System  of  Bedding 362 

The  Copper-matting  Blast-furnace 364 

Accessories  of  the  Blast-furnace 367 

Blast-furnace  Conditions 369 

Large  Copper-smelting  Blast-furnaces 369 

Regular  Operation  of  the  Copper  Blast-furnace 371 

Copper  Matte 373 


CONTENTS  xiii 


PAOE 


Copper-furnace  Slags 374 

Calculation  of  Charge  for  Matte  Smelting 375 

Pyrite  Matte  Smelting 377 

Reactions  in  Pyrite  Matte  Smelting 378 

Calculation  of  Charge  in  Pyrite  Smelting. 381 

Disposal  of  the  Slag 384 

Blast-furnace  vs.  Reverberatory  Smelting 385 


CHAPTER  XXX 

REVERBERATORY  SMELTING: 

The  Welsh  Process  of  Reverberatory  Smelting 387 

Smelting  Operations  by  the  Welsh  Process 388 

The  Direct  Process  of  Reverberatory  Smelting 390 

Large-scale  Reverberatory  Matte  Smelting 390 

The  Direct  Coal-fired  Furnace . . . . 391 

Furnaces  Fired  by  Pulverized  Coal 392 

The  Oil-fired  Furnaces 395 

Operation  of  a  Large  Reverberatory  Furnace 397 

Reactions  and  Calculation  of  the  Charge 397 


CHAPTER  XXXI 
CONVERTING  COPPER-MATTE: 

The  Copper  Converter 400 

The  Converter  Lining 402 

Operation  of  the  Basic  Converter 403 

Chemical  Reactions  of  the  Converter 404 

Blast-furnace  Smelting  and  Converting  Plant 406 

Electrostatic  Recovery  of  Copper  Blast-furnace  and  Converter  Dust 410 

Works  of  the  International  Smelting  Co 410 

Costs  of  a  Proposed  Plant  and  Operation 414 


CHAPTER  XXXII 

THE  HYDROMETALLURGY  OF  COPPER: 

Principles  of  the  Hydrometallurgy  of  Copper 417 

Extraction  of  Copper  by  Natural  or  Weathering  Methods 418 

The  Rio  Tinto  Process 418 

The  Shannon  Copper  Company  Process 421 

Extraction  of  Copper  as  a  Chloride f 422 

The  Henderson  Process 423 

The  Laist  Process ". 426 

Sulphuric  Acid  Leaching 429 

The  Butte-Duluth  Process 430 

The  Ajo  Process 432 

Ammonia  Leaching 438 

Ammonia  Leaching  at  Kennicott 439 


xiv  CONTENTS 

CHAPTER  XXXIII 
REFINING  OF  BLISTER-COPPER: 

PAGE 

Copper  Refining 442 

Melting  and  Refining  Lake  Copper 444 

The  Making  of  Anodes  and  of  Commercial  Cathode  Copper 446 

CHAPTER  XXXIV 
ELECTROLYTIC  COPPER  REFINING: 

Electrolytic  Copper-refining  Plant 449 

Capital  Requirements 455 

Cost  of  Refinery  and  Operating  Costs 456 

Schedule  of  Copper  Ore  Prices 457 

PART  VI— LEAD 
CHAPTER  XXXV 

PROPERTIES  OF  LEAD  AND  ITS  ORES: 

Characteristics  of  Lead  Ores 461 

The  Smelting  of  Lead-bearing  Ores 463 

Smelting  on  the  Ore-hearth 463 

CHAPTER  XXXVI 
SILVER-LEAD  SMELTING: 

Silver-lead  Blast-furnace  Smelting 467 

Receiving,  Sampling  and  Bedding  of  Lead  Ores 467 

General  Arrangement  of  a  Smelting  Works 468 

Bedding  Ores  at  a  Custom  Works 469 

The  Silver-lead  Blast-furnace 471 

Open  and  Closed-top  Blast-furnaces .  475 

Operating  the  Blast-furnace 476 

Chemical  Reactions  and  Physical  Changes  of  the  Blast-furnace 479 

Slags  in  Silver-lead  Smelting 480 

Action  of  Various  Bases  in  Slags 481 

Fuel  in  Silver-lead  Smelting 483 

Calculation  of  a  Blast-furnace  Charge 484 

i 
CHAPTER  XXXVII 

PRODUCTS  OF  THE  BLAST-BURNACE  : 

Flue-dust 487 

The  Bag-house " 488 

Briquetting  Flue-dust 489 

Lead-copper  Matte 491 

Comparison  of  Matte-treatment  Methods 491 

Converting  of  Leady  Matte 492 

Selling  Price  of  Matte .  493 


CONTENTS  XV 

CHAPTER  XXXVIII 
PRODUCTION  OF  LEAD  ORES  AND  PRICES: 

PAGE 

Costs  of  Lead  Ores 494 

Ore  Prices;  Mississippi  Valley  Smelting  Works 495 

Colorado  and  Utah  Silver-lead  Smelteries 495 

Variation  in  Costs  Due  to  Output,  etc 497 

CHAPTER  XXXIX 

REFINING  OF  LEAD  AND  BASE  BULLION: 

Refining  Base-bullion 498 

v  The  Refinery 499 

Softening  Base-bullion 500 

The  Parkes  Process 503 

Treatment  of  the  Rich  Lead 507 

The  Pattinson  Process. 510 

Cost  of  Refining  Base-bullion 511 

Selling  Price  of  Base-bullion 511 

The  Betts  Process  for  the  Electrolytic  Refining  of  Lead 511 

Smeltery  and  Refinery  for  Silver-lead  Ores 513 

PART  VII— ZINC 

CHAPTER  XL 
ZINC  AND  ITS  ORES: 

Properties  of  Zinc 517 

Zinc  Ores 517 

CHAPTER  XLI 

ROASTING  ZINC  ORES: 

Reduction  of  Ores  of  Zinc 519 

Roasting  Blende 519 

Chemistry  of  Roasting  Zinc  Ores 519 

Roasting  Furnaces 520 

The  Wedge  Mechanical  Blende-Roasting  Furnace 521 

The  Hegeler  Furnace 521 

Various  Furnaces 525 

The  Merton  Furnace 525 

The  Ridge  Furnace 525 

Sulphuric  Acid 527 

CHAPTER  XLII 

SMELTING  OF  ZINC  ORES: 

Smelting  or  Distillation  of  Roasted  Zinc  Ores 528 

Operating  the  Furnace 534 

Manufacture  of  Retorts  and  Condensers 536 

Loss  in  the  Process 537 

Cost  of  Smelting 538 

Price  of  Zinc  Ores  and  Spelter  in  1919 539 


xvi  CONTENTS 

CHAPTER  XLIII 
ZINC  REFINING: 

PAGE 

Grades  of  Spelter : 540 

The  De  Saulles  Redistillation  Method 542 

Refining  Spelter  without  Redistillation 542 

Electrolytic  Zinc 542 

PART    VIII— PLANT    AND    EQUIPMENT    AND    THEIR    COSTS 

CHAPTER  XLIV 

LOCATION,  EQUIPMENT  AND  ERECTION: 

Location  of  Works 547 

Nature  of  the  Site  to  be  Chosen 548 

Construction  of  Plant 550 

CHAPTER  XLV 

ACCESSORY  EQUIPMENT  OF  PLANT: 

Intermittent  Handling  of  Materials 551 

Industrial  Locomotives 551 

Industrial  Cars  and  Hoists 553 

Grabs  and  Excavators 555 

Continuous  Handling  of  Materials 557 

CHAPTER  XLVI 
ORE  STORAGE  AND  SUPPLY: 

Provision  for  Supply 563 

Feeders 563 

Pumps  and  Elevators 566 

CHAPTER  XLVII 
COST  OF  PLANTS: 

Cost  of  Plant • 569 

Cost  of  Metallurgical  Plants 571 

Unit  Construction  Costs  in  1914 . 572 

Composite  Costs 575 

% 

PART  IX— THE  BUSINESS  OF  METALLURGY 
CHAPTER  XLVIII 

THE  GENERAL  ECONOMIC  SITUATION: 

Distribution  of  Wealth . 579 

Economics  of  Engineering 579 

The  Labor  Situation 580 

Financial  Crises  in  the  U.  S 583 

Association  of  Employers 583 


CONTENTS 

CHAPTER  XLIX 
ORGANIZATION  AND  OPERATING: 

Organization  of  a  Metallurgical  Company 5g4 

The  Administrative  Department 5^4 

The  Operating  Department 5^5 

Rules  of  Works ggy 

Plant  Operation 5g7 

Morale  of  Inside-men : ggg 

Modes  of  Payment 5gg 

Capital  Requirements 59^ 

The  Accounting  Department 592 

Administration  and  General  Charges 594 

Typical  Operating  Department 594 

The  Purchasing  and  Selling  Departments 595 

CHAPTER  L 
PROFITS  AND  COSTS: 

Profits 597 

Custom  Smelteries 59g 


ERRATA 

page      8 — Figure  reference  14,  3d  line  from  top,  should  read  Figure  4. 

page    77 — Figure  reference  7,  7th  line  from  top,  should  read  Figure  14. 

page     78 — Figure  reference  122,  5th  line  from  bottom,  should  read  Figure  212. 

page    84 — Figure  reference  181A,  7th  line  from  top,  should  read  Figure  278. 

page    93 — Figure  reference  31,  llth  line  from  bottom,  should  read  Figure  70. 

page  130 — Figure  references  29  and  30,  last  line  should  read  Figures  60  and  61. 

page  135 — Figure  reference  70,  7th  line  from  bottom,  should  read  Figure  71. 

page  260 — Figure  reference  117,  1st  line,  should  read  Figure  116A. 

page  322 — Figure  reference  104,  12th  line  from  top,  should  read  Figure  169. 

page  388 — Figure  reference  29,  15th  line  from  top,  should  read  Figure  63. 

page  405 — Figure  131,  should  read  Figure  221. 

page  421 — Figure  221,  should  read  Figure  230. 

page  446 — Figure  reference  147,  1st  line,  should  read  Figure  242. 

page  465 — Figure  251,  omitted. 

page  478 — Figure  reference  155,  14th  line  from  bottom,  should  read  Figure  263. 

page  523 — Figure  283,  should  read  Figure  282. 

page  525 — Figure  reference  384,  9th  line  from  top,  should  read  Figure  284. 

page  563 — Figure  references  15A,  40,  45,  98,  7th  line  from  botton,  should  read 
Figures  28.  76,  83  and  134. 


. 


PART  I 
GENERAL  METALLURGY 


CHAPTER  I 

ORES  AND  METALS 

DEFINITION  AND  CLASSIFICATION  OF  ORES 

Definition. — An  ore  from  the  standpoint  of  the  metallurgist  may  be 
denned  as  a  mineral  aggregate  containing  metal,  or  metals,  in  sufficient 
quantity  to  make  their  extraction  commercially  profitable.  Minerals  or 
rocks  containing  15  to  30  per  cent  iron  would  not  be  called  iron  ore,  nor 
would  we  call  a  rock  containing  2  to  3  oz.  silver  per  ton  a  silver  ore.  On 
the  other  hand,  the  rock  of  the  Treadwell  mine,  on  Douglas  Island,  Alaska, 
carrying  $2.50  to  $3  in  gold,  is  called  a  gold  ore  because  it  can  be  worked 
at  a  profit.  In  general,  ores  are  named  from  their  dominant  metal  (as 
lead,  copper,  or  silver),  though  they  may  contain  other  metals.  Thus  a 
lead  ore  may  contain  silver  and  gold;  a  copper  ore,  besides  copper,  may 
contain  silver,  gold,  and  even  lead.  The  appearance  of  an  ore  may  indi- 
cate whether  it  carries  lead,  copper,  iron,  or  zinc,  but  gold  and  silver  min- 
erals are  not  always  visible,  and  the  proper  way  to  determine  their  presence 
is  by  assay. 

Straight  or  simple  ores  contain  in  the  main  but  one  kind  of  metal,  such 
as  gold,  silver,  copper,  or  lead.  Straight  silver,  or  free-milling  silver  ores, 
are  free  from  lead  and  copper,  and  may  be  treated  by  amalgamation. 
Straight  gold  ores,  also  free-milling,  are  those  containing  the  gold  in 
metallic  form  and  amenable  to  amalgamation.  Straight  or  plain  lead, 
zinc,  or  copper  ores  do  not  contain  gold  or  silver  in  quantity  sufficient  to 
pay  to  separate  the  precious  metals  from  the  base  metal.  As  an  example, 
a  lead  ore  containing  4  oz.  silver  per  ton  would  not  ordinarily  meet  the 
cost  of  extracting  the  silver.  Blister  copper  may  contain  as  much  as  12  oz. 
silver  per  ton  and  yet  not  pay  the  charge  for  electrolytic  refining  for  its 
recovery.  An  ore  containing  little  lead,  say  less  than  5  per  cent,  is  desig- 
nated a  jlry  ore.  _Such  ore  is  often  silicious,  but  possesses  commercial 
value  because  it  contains  gold  and  silver.  Ores  carrying  more  than  5 
to  10  per  cent  lead  may  be  profitably  treated  for  their  lead  alone.  Copper 
ores  also  frequently  contain  gold  and  silver. 

Mixed  ores,  or  those  containing  two  or  more  kinds  of  metal,  are  com- 
mon, such  as  silver-gold,  silver-gold-lead,  or  lead-zinc-copper-silver. 
When  such  ores  contain  both  copper  and  lead  it  is  puzzling  at  times  to 

3 


ORES  AND  METALS 

know  how  to  designate  them.  In  doubtful  cases  smelting  companies 
have  purchased  them  either  on  the  basis  of  their  lead  or  copper  content 
under  the  plea  that,  in  extracting  one  of  these  metals,  the  other  is  lost  or 
wasted.  Lead-silver  or  lead-silver-gold  ores  are  those  which  carry  lead 
in  such  quantity  that  when  the  lead  is  recovered  from  them  by  smelting, 
the  precious  metals  taken  up  by  it  can  be  later  easily  removed  from  the 
lead.  Copper-silver,  copper-silver-gold,  or  copper-gold  ores,  when  smelted, 
yield  their  copper,  and  this,  like  lead,  takes  up  the  precious  metals. 

Base-metal  Ores. — Lead  and  copper  ores  often  contain  zinc,  antimony, 
arsenic,  tellurium,  or  bismuth  as  impurities.  These,  in  the  process  of 
reduction,  alloy  with  the  principal  metal  to  its  commercial  detriment,  and 
require  expensive  after-treatment  to  remove  them.  While  a  free-milling 
ore  permits  the  extraction  of  most  of  its  gold  or  silver  by  simple  processes 
of  grinding  and  amalgamation,  a  refractory  or  rebellious  ore  requires  pre- 
liminary treatment  by  roasting  before  it  can  be  amalgamated;  otherwise 
it  must  be  smelted.  Even  smelting  ores  may  present  difficulties  of  treat- 
ment that  would  cause  them  to  be  called  rebellious.  A  docile  ore,  on  the 
contrary,  is  one  that  may  be  easily  treated.  Gold  and  silver  ores  con- 
taining arensic  or  antimony  may  be  cited  as  examples  of  refractory  ores. 

METHODS  OF  TREATMENT 

These  may  be  divided  broadly  into  milling  or  smelting  plants.  Mills 
treat  gold  and  silver  ores,  according  to  their  character,  by  concentra- 
tion (including  flotation),  amalgamation,  chlorination,  cyaniding,  or  by 
combinations  of  these  methods.  Thus,  the  North  Star  Mine,  Grass 
Valley,  Cal.,  treats  a  gold  and  silver  quartz  ore  (carrying  sulphides)  by 
amalgamation,  concentration  and  cyaniding  of  the  concentrates  and 
tailings.  The  Tonopah-Belmont,  a  medium  hard  quartz  silver  ore 
carrying  sulphides,  uses  concentration  and  cyaniding.  Another  mill,  the 
Liberty  Bell,  Telluride,  Colo.,  having  a  soft  quartz  silver  ore, 
subjects  it  to  amalgamation,  concentration,  and  cyaniding  of  the 
tailings.  Smelting  ores  may  be  basic,  silicious,  dry,  coppery,  or  leady; 
while  milling  ores  may  be  talcose,  quartzose,  raw,  roasting,  earthy, 
argillaceous,  light,  heavy,  or  base,  all  of  which  characteristics  modify  the 
mode  of  treatment.  Among  the  iron  ores  we  may  have  Bessemer  ores  or 
those  containing  not  more  than  0.045  per  cent  phosphorus,  and  non-Bes- 
semer ores,  or  those  so  high  in  phosphorus  that  the  pig  iron  made  from 
them  needs  subsequent  treatment  in  the  basic  open-hearth  furnace  to 
remove  it. 

An  ore  consists  not  only  of  the  species  of  metallic  compound  from 
which  it  is  named,  but  also  of  gangue  or  waste  matter.  This  may  often 
be  its  principal  constituent,  and  may  be  earthy,  silicious,  argillaceous, 


METHODS  OF  TREATMENT  5 

talcose,  or  limy,  and  the  ore  may  be  composed  largely  of  the  lighter 
gangue  with  comparatively  small  quantities  of  the  valuable  metals  scat- 
tered or  disseminated  through  it.  When,  as  is  often  the  case,  the  metal 
is  the  heavy  part  of  the  ore,  and  the  lighter  part  is  the  gangue,  the  ore 
may  be  concentrated  or  dressed  with  a  view  to  removing  this  gangue. 
An  ore  capable  of  being  thus  treated  is  called  a  concentrating  ore,  and  the 
valuable  heavy  part  obtained  from  it  is  called  a  "  concentrate." 

We  may  also  divide  ores  into  sulphide  and  oxidized.     As  a  matter  of  r 
fact,  these  merge  into  one  another,  and  it  is  often  difficult  to  decide  to 
which  class  to  assign  a  given  ore.     Carbonates  are  placed    among    the 
oxidized  ores,  since,  in  smelting,  the  carbon  dioxide  is  readily  driven  off, 
leaving  the  oxide  of  the  metal. 

Grading  Ore. — Miners  often  find  it  profitable  to  sort  their  ore  into 
different  grades,  such  as  shipping  or  smelting,  and  into  milling  or  con- 
centrating ore,  according  to  the  after-treatment  they  purpose  to  give  it. 
This  matter  is  often  an  important  one  for  the  metallurgist  to  consider  in 
deciding  upon  the  treatment  of  ore,  as,  for  example,  in  the  case  of  a  mixed 
silver  ore. 

CLASSIFICATION  OF  METALLURGICAL  OPERATIONS 

These  may  be  roughly  divided  into  two,  viz.,  milling  and  smelting. 

Gold  and  silver  ores  are  commonly  treated  in  mills,  though  they  may 
also  be  smelted. 

Iron,  lead,  zinc,  and  copper  ores  are  commonly  smelted,  though  the 
last  three  can  be  treated  by  hydrometallurgical  methods  in  which  the 
metal  is  brought  into  solution  and  later  precipitated  from  the  solution. 
We  speak  then  of  smelting  being  a  pyrometallurgical  process,  while  hydro- 
metallurgy  relates  to  the  extraction  of  the  metal  by  aqueous  solutions. 
In  smelting,  the  ore  is  roasted  if  necessary,  smelted  in  a  smelting  furnace, 
and  the  product,  the  metal  still  containing  impurities,  refined  to  put  it  in 
marketable  form.  Gold  and  silver  have  been  successfully  recovered 
from  free  milling  ores  by  crushing  and  amalgamation.  However,  most 
ores  cannot  be  so  easily  treated.  The  steps  of  milling  practice  then  are 
(1)  crushing  and  grinding,  (2)  solution,  (3)  filtration,  (4)  precipitation, 
(5)  refining,  as  given  in  later  chapters  of  this  book. 

PRINCIPLES  RELATING  TO  THE  REFINING  OF  METALS 

It  is  found  by  analysis  that  the  separation  of  a  metal  from  other  metals 
or  from  contained  impurities  is  seldom  complete.  It  is  difficult  and  com- 
mercially impracticable  to  obtain  metals  entirely  pure,  so  that  those 
that  come  on  the  market  still  contain  small  amounts  of  impurity.  Metals 
thus  prepared  are  graded  according  to  quality,  and  command  prices 


ORES  AND  METALS 


TABLE  I.— PHYSICAL  CONSTANTS  (METALS  AND  GASES) 


Sym- 

Atomic 

Specific 
Gravity, 

SPECIFK 
WATEE 

3   HEAT, 
=  1.00.    . 

Melting 

Latent 
Heat  of 

Tensile  Strength 

Metal. 

bol. 

Weight. 

Water 
=  1.00 

30°  C. 

Melting 
Point. 

Point 

°c. 

Fusion, 
Calories. 

Lbs.  per  Sq.  In. 

Aluminium  .... 

Al 

27 

2.56 

0.167 

0.308 

659 

100     { 

12,590  cast 
19,290  rolled 

Antimony  

Sb 

120 

6.71 

0.048 

0.054 

630 

40 

1,000 

Arsenic  

As 

75 

5.67 

0.076 

850 

Bismuth  

Bi 

137 

9.8 

0.031 

271 

12 

3,000 

Cadmium  

Cd 

208 

8.6 

0.054 

0.062 

321 

13 

Calcium  

Ca 

40 

1.57 

0.170 

810 

Carbon 

c 

12 

3650 

Chromium  

Cr 

52 

6.8 

0.104 

1510 

Cobalt  

Co 

59 

8.5 

0.106 

0  204 

1490 

68 

75,000 

Copper  

Cu 

63 

8.8 

0.086 

0.118 

1083 

•  { 

34,000  bolts 
60,000  wire 

Gold  

Au 

196 

19.3 

0.032 

1063 

16    { 

20,000  cast 

I 

37,000  wire 

Iridium  

Ir 

192 

22.4 

0.030 

0.04 

2300 

Iron  . 

Fe 

56 

7.86 

0.116 

0.162 

1520 

69    / 

55,000  rolled 

I 

48,000  cast 

Lead  

Pb 

206 

11.4 

0.030 

0.034 

328 

4    ( 

2,650  cast 

\ 

1,650  pipe 

Magnesium.  .  .  . 

Mg 

24 

1.74 

0.246 

651 

Manganese.  .  .  . 

Mn 

55 

8.00 

0.122 

1225 

Mercury 

Hg 

200 

13.6 

0.033 

0.032 

—39 

3 

Nickel  

o 

Ni 

59 

6.7 

0.109 

0.161 

1452 

68 

54,000 

Osmium  

Os 

195 

22.5 

0.031 

2700 

Palladium  

Pd 

106 

11.5 

0.059 



1549 

36 

50,000 

Phosphorus.  .  .  . 

P 

31 

21.5 

Platinum  

Pt 

194 

21.5 

f  0.032 

0.046 

1755 

27    { 

45,000  cast 

I 

56,000  wire 

Potassium  

K 

39 

0.87 

0.166 

62 

Silver  

Ag 

108 

10.5 

0.055 

0.076 

960 

24 

41,000 

Sodium  

o 

Na 

23 

0.97 

0.293 

98 

Tin  

Sn 

117 

7.3 

0.055 

0.059 

232 

14    ( 

4,600  cast 

I 

5,800  drawn 

Tungsten  

W 

184 

19.1 

0.034 

3000 

Zinc  .... 

Zn 

65 

7.1 

0.093 

0.112 

419 

23 

«,vw 

GASES 


Chlorine  
Fluorine 

Cl 

Fl 

35 
19 

Hydrogen  
Nitrogen  

H 

N 

1 
14 

0.00009 
0.0125 

Oxygen  

O 

16 

0.0143 

- 

MOLDING  AND  CASTING  METALS  7 

according  to  the  grade.  Thus  Lake  copper  commands  the  highest  price  of 
any  copper  because  of  its  purity  and  toughness,  while  electrolytic  copper 
sells  at  one-half  cent  less  per  pound. 

In  silver-lead  smelting  practice  the  slag,  no  matter  how  thoroughly 
settled  and  separated  from  the  matte,  still  contains  0.2  to  0.3  oz.  silver 
per  ton  and  0.3  to  0.4  per  cent  lead.  In  copper  refining,  in  the  reverbera- 
tory  furnace,  arsenic,  antimony,  and  bismuth,  occurring  in  the  crude  or 
blister  copper,  are  retained  as  traces  after  refining,  and  where  the  blister 
copper  is  impure,  no  high-grade  product  can  be  expected.  In  the  sep- 
aration and  deposition  of  copper  by  electrolysis  at  low  current-density, 
the  copper  is  of  high  grade  even  though  impurities  are  in  solution  in  the 
electrolyte;  nevertheless,  traces  of  impurity  find  their  way  into  the  cathode 
copper,  though  to  less  extent  than  by  any  other  system  of  refining. 

In  the  refining  of  pig  iron  to  make  steel,  in  order  to  obtain  satisfactory 
quality,  impurities  must  be  removed  until  less  than  0.10  per  cent  phos- 
phorus and  0.05  per  cent  sulphur  are  present,  otherwise  the  steel  lacks 
toughness  and  tenacity. 

MOLDING  AND  CASTING  METALS 

Metals  undergoing  treatment  are  finally  brought  to  the  metallic  state, 
and  are  commonly  cast  into  ingots  or  bars  for  sale.  Sometimes  metals 
may  be  finally  granulated,  as  zinc  for  cyaniding  or  lead  for  test-lead  in 
assaying.  Or  again,  the  molten  metal  may  be  poured  into  water,  pro- 
ducing coarse  flattened  granulations  where  it  is  desired  to  quickly  dissolve 
the  metal,  as  in  the  parting  of  precious  metals.  -  ^ 

But  generally  metals  are  cast  into  bars,  ingots  or  commercial  shapes 
as  desired  by  the  customer,  who  wishes  to  subject  such  bars  to  further 
treatment.  These  shapes  vary  according  to  the  metal  and  are  cast  in 
molds  either  by  hand  or  by  casting  machines  as  follows : 

Gold. — This  is  cast  of  all  sizes  according  to  the  quantity  treated,  or 
to  the  desire  of  the  customer,  from  the  size  of  the  finger  upward,  to  be 
rolled  into  sheets,  to  be  drawn  into  wire  or  to  be  sold  to  the  mint.  Gold 
for  the  mint  is  then  remelted  in  plumbago  crucibles  of  suitable  size,  assayed 
and  paid  for,  then  granulated  for  parting. 

Silver  is  commonly  cast  in  bars  of  1000  oz.  (about  80  lb.),  a  con- 
venient size  for  handling,  and  hard  to  steal.  When  ready  to  tap  from  the 
test  of  a  cupelling  furnace  it  is  run  into  ingot  molds  standing  on  a  car- 
riage beneath,  then  pushed  along  as  the  ingots  fill.  In  similar  quantity 
the  metal  may  be  remelted  in  a  plumbago  crucible,  then  lifted  out  with 
basket  tongs  and  poured  directly  into  the  molds,  as  in  silver-mill  practice. 
Where  the  quantity  of  metal  is  large,  as  in  the  treatment  of  precipitate, 
the  melt  may  be  poured  into  a  crucible  and  thence  into  molds,  as  shown 


8 


ORES  AND  METALS 


in  Fig.'  2.     The  tilting  furnace  is  oil-fired   and  is   shown   in   skimming 
and  pouring  position. 


FIG.  1. — Plumbago  Crucibles. 


FIG.  2. — Conical  Mold. 


Fig.  14  is  a  mold  such  as  is  used  for  silver  bars  11  in.  long  by  4J 
in.  by  4J  in.,  to  hold  1000  oz.  or  70  Ib.  of  the  metal. 

Iron  and  Steel. — Pig  iron 
from  the  blast-furnace  was  for- 
merly cast  upon  the  sand  floor 
of  the  cast-house.  The  furnace 
was  placed  centrally  at  one  end 
of  the  house,  and  the  floor  sloped 
away  at  even  grade  from  the 
metal  tap-hole.  There  was  a 
central  channel  or  runway  made 
in  the  sand  to  the  end  of  the 
house.  Branching  either  way  at 
right  angles  from  this  channels 
were  made  in  the  sand,  and  by 
means  of  wood  molds,  cavities 
for  the  pigs.  Thus  the  branches 
were  fancifully  called  the  sows, 
the  cavities  the  pigs.  When  iron 


FIG.  3.— Tilting  Furnace. 


FIG.  4. — Cast-iron  Mold. 


FIG.  5. — Bosh  and  Molds. 


was  tapped  the  flow  was  down  the  main  channel.     To  divert  the  flow, 
gates  (flat  cast-iron  plates)  were  so  set  across  the  channel  as  to  change 


MOLDING  AND  CASTING  METALS  9 

it  to  any  branch,  filling  the  cavities  forming  the  pigs.  When  they  were 
filled  the  gate  was  set  at  the  next  branch  the  main  run  was  opened,  and 
the  filled  side  branch  cut  off,  all  by  means  of  the  gates.  After  cooling  the 
pigs  were  broken  off  by  prying  them  up  with  a  bar,  and  removed  to  the 
cars  standing  upon  tracks  alongside  the  cast-house. 

The  pigs  had  some  sand  sticking  to  them,  and  this  was  one  reason  that 
mechanical  casting  was  adopted,  as  in  the  Heyl  and  Pattison  machine, 
Fig.  162. 

Pig  iron,  as  we  know,  upon  remelting  in  a  cupola  furnace,  makes  the 
most  intricate  castings.  The  same  is  true  of  steel  castings,  now  success- 
fully melted  in  the  converter  or  in  the  electric  furnace,  treated  by  the 
addition  of  ferro-alloys  to  make  a  quiet  melt,  and  producing  strong  castings 
for  special  purposes. 

Copper. — This  metal  after  refining  is  cast  into  special  forms  for  the 
market,  in  small  furnaces  by  hand.  The  skimmed  metal  is  dipped  from 
the  furnace,  using  a  dipping  ladle  having  a  bowl  9  in.  diameter  and  holding 
25  Ib.  In  casting,  a  water-bosh  is  used,  see  Fig.  5.  On  the  edge  of 
this  is  hinged  a  number  of  molds  which  are  successively  filled  by  three 
or  four  men  who  dip  from  the  furnace  and  fill  them.  As  soon  as  these 
get  solid,  the  mold  is  turned  over,  the  ingot  falling  into  the  water. 
From  the  water  it  is  picked  out  by  means  of  tongs  and  is  ready  for 
market.  The  usual  shapes  are  ingots,  ingot  bars,  wire  bars,  and  cakes, 
which  weigh  respectively  17  Ib.,  35  Ib.,  85  to  250  Ib.,  and  rectangular  cakes 
14  in.  square  or  14  by  17  in.,  to  roll  into  plates. 

In  place  of  the  hand  ladles,  often  "  bull  ladles  "  are  used.  These, 
having  long  handles,  suspended  near  the  center  of  balance  by  chain  to  an 
overhead  trolley  rail,  will  dip  up  100  Ib.  at  a  time.  They  greatly  expedite 
the  work  of  dipping.  This,  with  an  ordinary  refining  furnace,  will  take 
four  or  five  hours. 

In  place  of  hand  dipping,  one  of  the  mechanical  casting  machines, 
such  as  the  endless  mold  machine,  Fig.  242,  or  the  Walker  casting  machine, 
Fig.  245,  is  employed.  Indeed,  for  one  of  the  large  furnaces,  hand  dipping 
would  be  too  slow.  As  seen  in  Fig.  245  the  machines  are  supplied  by  a 
ladle  from  the  converter,  and  blister  copper  in  ingots  of  250  to  400  Ib.  are 
cast. 

In  place  of  a  casting  machine  a  series  of  molds  are  often  employed, 
enough  of  them  to  take  care  of  a  ladle  full  of  blister  copper.  When  a 
works  desires  to  produce  a  finished  cathode,  this  is  done  by  collecting  the 
converted  product  in  a  tilting  furnace,  resembling  the  tilting  open-hearth 
furnace  of  steel  practice.  It  is  here  poled,  making  a  smooth  ingot,  or  even 
an  anode  suitable  for  use  in  the  electrolytic  vat. 

Lead. — The  older  way  of  casting  base-bullion  was  to  let  the  molten 
metal  run  into  a  cooler,  a  basin  that  would  hold  1000  Ib.  The  dross  was 


10  ORES  AND  METALS 

skimmed  from  this,  and  the  base-bullion  dipped  by  ladle  into  molds 
holding  85  to  100  lb.,  then  shipped  to  a  refinery.  The  present  practice  is 
to  receive  the  molten  metal  into  a  two-wheeled  pot.  It  is  then  taken  to  the 
drossing  kettle,  where  the  dross  is  removed  as  described  under  lead  refining. 
The  metal,  now  free  from  dross,  is  shipped  away  to  the  refinery.  The 
dross  is  returned  to  the  blast  furnace.  It  is  a  coppery  dross  still  containing 
lead. 

Under  head  of  Refining  we  give  the  method  used  for  molding  market 
lead.  The  lead  when  solid  is  removed  from  the  molds  by  hand.  In  some 
cases  an  endless  mold  machine  is  used. 

Zinc. — This  is  tapped  from  a  horizontal  row  of  condensers  into  a  ladle 
suspended  by  chain  blocks  to  a  trolley  rail.  The  collected  metal  is  skimmed 
from  dross  and  poured  into  molds.  In  remelting  spelter,  a  charcoal  cover 
should  be  used,  since  hot  molten  zinc  easily  drosses.  It  is  for  this  reason 
that  in  brass  making  the  copper  is  first  melted  and  the  zinc  added  at  the 
last  moment  before  casting. 


CHAPTER  II 
FUELS 

A  fuel  may  be  defined  as  a  solid,  liquid,  or  gaseous  substance  that  can 
be  burned  for  the  production  of  heat  for  economic  purposes.  Fuels  can 
be  divided  into  two  classes:  natural  and  artificial.  Coal  is  a  natural  fuel; 
coke  an  artificial  one.  The  natural  fuels  include  "  solid  fuel  "  like  wood 
or  coal,  the  mineral  oils  and  natural  gas.  The  solid  natural  fuels  are 
believed  to  be  of  vegetable  origin.  They  are  substances  in  some  measure 
altered  from  their  original  condition  by  heat  and  pressure,  and  range  from 
wood  through  peat,  lignite,  bituminous  or  soft  coal,  anthracite  or  hard  coal 
to  graphite  at  the  extreme.  Artificial  fuels  may  be  divided  into  the 
"  solid  prepared  fuels  "  and  "  fuel 

3  3 

|         a     »      -g     3  | 

i     I    II 


gas."      The  solid  fuels  are  coke 


and  charcoal. 

The  relation  of  the  natural 
carbonaceous  substances  is  shown 
in  Fig.  7.  Here  in  a  general  way 
is  illustrated  the  chemical  and 
physical  changes  that  occur  in 
the  formation  of  coal  from  its 
organic  constituents  in  plant- 
tissue.  These  changes  result, 
finally,  under  the  action  of  pres- 
sure and  high  temperature,  in 
graphitic  carbon,  and  begin  by 
action  upon  wood,  leaves,  and 
root-fibers  (turf  or  peat),  passing 
through  lignites  or  freshly  formed 
coal,  often  brown  in  color,  thence 


1 


100 


FIG.  7. — Table  Showing  Genesis  of  Natural 
Fuels. 


to  bituminous  coal  formed  during  recorded  geological  time,  retaining  the 
volatile  constituents,  to  anthracite  where  the  volatile  constituents  are 
mostly  eliminated  by  the  heat  and  immense  pressure  and  finally  result  in 
graphite  where  distillation  completes  the  work. 


11 


12  FUELS 

THE  NATURAL  SOLID  FUELS 

Classification. — A  convenient  division  of  the  standard  types  of  such 
fuels  may  be  made  into  first,  those  of  compact  texture,  and  second,  those 
of  woody,  fibrous,  or  earthy  texture. 

A  more  exact  and  convenient  classification,  beginning  with  the  most 
compact,  the  highest  in  "  rank,"  is  thus  given: 

(a)  Graphite,  native  coke,  and  anthracites,  these  burning  with  a 
non-luminous  flame. 

(6)  Bituminous  coal  or  bitumen,  burning  with  a  luminous  flame. 

(c)  Lignites,  peat,  and  wood,  fuels  having  a  woody,  fibrous,  or 
earthy  structure,  burning  with  a  luminous  flame. 

The  impurities  of  coal  are  ash,  sulphur,  and  to  a  lesser  extent  nitrogen. 
Of  a  given  type  the  standard  may  be  given  at  6  per  cent  ash,  and  1  per  cent 
sulphur.  In  nitrogen  0.75  per  cent  for  anthracite;  1.5  per  cent  in  the 
intermediate  types,  decreasing  to  0.75  per  cent  in  the  lignites.  These 
impurities  cause  a  variation  in  "  grade." 

The  rank  of  a  coal,  in  changing  from  peat  to  graphite  (see  Table  II), 
shows  a  progressive  elimination  of  moisture  and  volatile  matter  and  a  cor- 
responding increase  in  the  proportion  of  fixed  carbon  and  ash.  Thus,  a  typi- 
cal fresh  peat  would  contain  91  per  cent  moisture,  6  per  cent  volatile  matter, 
2  per  cent  fixed  carbon,  and  0.3  per  cent  ash.  A  typical  lignite,  assumed  to 
have  been  derived  from  the  peat,  contains  43  per  cent  moisture,  26  per  cent 
volatile  matter,  27  per  cent  fixed  carbon,  and  25  per  cent  ash.  If  from 
lignite  we  go  on  up  through  the  list,  we  find  that  while  the  amount  of 
moisture  in  the  coal  steadily  decreases  the  percentage  of  volatile  matter 
keeps  about  even  with  that  of  the  fixed  carbon  in  all  the  lower-rank  coal 
until  the  moisture  reaches  a  stable  minimum  beyond  which  the  percentage 
of  volatile  matter  is  rapidly  reduced.  Thus  we  find  the  ratio  of  volatile 
matter  plus  H^O  to  the  fixed  carbon  is  one-to-one  in  the  lower-rank  coals; 
it  has  risen  to  one-to-four  in  the  best  bituminous,  and  to  one  in  seven  or 
more  in  the  anthracite. 

The  Anthracites. — These  have  a  fuel  ratio  of  one  to  seven  and  burn  with 
a  non-luminous  flame.  They  are  conveniently  grouped  into  the  "  hard," 
having  a  conchoidal  fracture,  high  specific  gravity  and  sub-metallic 
luster;  and  the  soft  with  a  semi-cubic  fracture  and  low  specific  gravity. 

The  Bitumites. — These  include  the. bituminous  coals  of  the  carbon- 
iferous age  and  the  sub-bituminous  coals,  or  those  of  the  post-carbon- 
iferous, both  having  a  fuel  ratio  less  than  one  to  seven  and  burning  with  a 
luminous  flame.  This  flame  indicates  the  presence  of  hydrocarbons  in 
the  volatile  constituents,  and  so  that  the  coal  is  bituminous. 

The  coals  of  the  carboniferous  age  are  divided  into  (1)  the  so-called 
smokeless  Virginia  coals,  which  includes  those  having  a  short  and  those 
having  a  medium  flame;  (2)  the  coking  or  steam  coals,  having  a  long  flame; 


COMPOSITION  OF  COALS 


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14  FUELS 

(3)  the  non-coking  or  "  household  "  coals,  having  a  woody  texture  and 
cubic  fracture;  household  coals  also  having  the  same  texture,  but  a  con- 
choidal  fracture. 

The  coals  of  the  post-carboniferous  age  are  divided  into  (1)  the  weather- 
resisting  or  Montana  coals,  which  may  be  stocked  or  shipped  long  dis- 
tances, and  (2)  the  non-weather-resisting  or  semi-bituminous,  which  exposed 
to  alternate  wetting  and  drying,  break  down  and  lose  their  shape  in  a  month's 
time.  In  this  latter  group  is  found  New  Mexico  coal,  having  a  fuel-value 
of  more  than  7750  calories,  and  the  Wyoming  coal  with  less  than  that. 

The  lignites  are  high  in  moisture,  the  black  lignites  or  sub-bituminous 
coals  containing  much  less  than  30  per  cent,  the  brown  coals  more  than 
30  per  cent  moisture.  They  have  a  woody  structure  and  a  luminous  flame. 

In  the  following  table  of  coals  of  the  United  States,  we  give  the  ratio 
obtained  by  dividing  the  fixed  carbon,  as  found  in  the  proximate  analyses, 
by  the  combined  percentages  of  volatile  combustible  and  moisture.  The 
calorific  value  of  the  coal  and  its  analysis,  both  proximate  and  ultimate, 
are  also  given.  Of  the  two,  the  former  is  the  one  generally  thought  of. 
A  coal  of  standard  type  is  that  used. 

Wood. — When,  freshly  cut,  wood  contains  40  per  cent  moisture,  and  in 
this  condition  is  difficult  to  burn  alone;  but  where  this  can  be  done,  it 
develops  2300  pound-calories  per  pound.  Split  into  cord-wood,  piled,  and 
dried  for  several  months,  wood  contains  20  per  cent  moisture  and  40  per 
cent  carbon.  Its  calorific  value  thereby  increases  to  3600  calories.  Such 
wood  is  classed  as  hard  when  its  specific  gravity  is  more  than  0.55;  below 
this  it  is  called  soft.  While  the  calorific  intensity  of  dry  wood  is  low,  its 
combustibility  is  great,  and  it  is  well  suited  for  use  in  reverberatory  roast- 
ing-furnaces,  since  the  volatile  constituents,  rapidly  escaping,  burn  grad- 
ually, and  make  an  extended  flame  along  the  hearth  of  the  furnace,  heating 
it  more  uniformly  than  could  a  flameless  fuel  like  anthracite  or  coke.  In 
the  outlying  districts  of  the  western  United  States,  where  the  metallurgist 
is  dependent  on  wood  for  generating  steam,  or  roasting  ore,  the  accumula- 
tion of  a  sufficient  supply  of  dry  wood  should  be  one  of  his  first  cares.  In 
this  his  forethought  is  well  rewarded.  He  should  purchase  wood  delivered 
and  corded  near  the  works;  and  in  measuring,  make  equitable  allowance 
for  short  dimensions  or  open  piling.  Cord  wood  should  "  cord  up  to 
70  per  cent  solid  wood. 

Coal  for  Roasting. — Both  lignites  and  the  regular  bituminous  coal  may 
be  used.  It  should  have  a  good  proportion  of  "  volatile  matter,"  so  that 
the  flame  may  be  long,  and  thus  distribute  itself  over  a  greater  area  of 
the  roasting  hearth.  A  short  flame  would  be  intense  near  the  fire-box, 
but  would  fail  farther  away.  To  make  the  flame  long  the  quantity  of  air 
is  so  regulated  that  it  mingles  slowly  with  the  escaping  volatile  gases,  and 
thus  it  is  carried  to  the  end  of  the  hearth. 


FUEL  OILS  15 

Coals  of  very  different  properties  may  appear  alike  if  represented  only 
by  proximate  analysis.  The  comparative  calorific  value  may  be  judged 
of  by  Berthier's  method.  This  consists  practically  in  the  operations  of  a 
lead  assay,  using  an  excess  of  litharge,  with  a  gram  of  the  fuel,  and  noting 
the  size  of  lead  button  reduced.  One  can  also  judge  a  good  deal  about  the 
character  of  the  coal  by  coking  it  in  a  covered  crucible  and  weighing  the 
coke  produced,  judging  the  character  by  the  appearance  of  the  product. 
The  proximate  analyses  (Table  II),  showing  the  different  kinds  of  coal, 
determine  to  which  class  any  given  kind  belongs. 

Graphite. — This  is  of  interest,  not  as  a  fuel,  but  as  a  refractory  material, 
particularly  when  combined  with  clay. 

Petroleum  or  Fuel  Oil. — This  is  the  most  concentrated  of  fuels,  and, 
when  the  cost  justifies,  can  be  used  not  only  for  generating  steam,  but  for 
roasting  and  melting.  It  will  be  found,  in  burning  fuel-oil  from  various 
localities,  that  the  calorific  power  is  much  the  same  for  the  different  kinds. 
Beaumont  (Texas)  oil  has  a  calorific  power  of  10,820  calories,  and  a  specific 
gravity  of  0.88  (7J  Ib.  per  gallon.)  Oil  can  be  burned  in  such  a  way  as  to 
give,  not  only  a  high  and  uniform  temperature,  but  also  the  oxidizing 
(roasting)  or  reducing  action  that  may  be  desired.  The  air  for  com- 
bustion is  best  preheated  as  well  as  the  oil,  and  it  will  be  found  advan- 
tageous to  inject  the  oils  under  a  high  steam  pressure.  A  mixture  of  light 
and  heavy  oils  should  not  be  used.  In  Russia,  where  it  has  been  employed 
in  open-hearth  steel-furnaces  of  10  to  15  tons  capacity,  oil  to  the  extent  of 
15  to  20  per  cent  of  the  weight  of  the  charge  has  been  used.  As  regards 
comparative  costs  at  the  Selby  Smelting  &  Lead  Works,  Vallejo  Junction, 
California,  it  was  found  (hat  the  saving  was  40  to  60  per  cent  with  oil  at 
$1.71  per  bbl.  (42  gal.)  and  coal  at  $6  per  ton.  A  suitable  control  of 
the  grade  of  the  matte  was  possible  by  the  regulation  of  the  flame. 

Natural  Gas. — In  Ohio,  Indiana,  and  Kansas,  particularly,  there  are 
districts  where  natural  gas  has  been  obtained  by  boring  for  it  as  for  oil. 
It  is  the  most  efficient  of  natural  fuels,  having  a  calorific  power  of  611  cal. 
per  cubic  ft.  or  27,862  Cal.  per  pound.  The  following  analysis  will  give 
an  idea  of  the  composition  of  Pennsylvania  natural  gas.  It  shows  that 
it  is  composed  chiefly  of  marsh-gas  and  hydrogen,  viz: 

Per  Cent, 
by  Volume. 

Carbon  dioxide  (CO2) 0.8 

Carbon  monoxide  (CO) 1.0 

Oxygen  (O2) 1 . 1 

Ethylene  (CzII*) .,-. .........  1.0 

Ethane  (CzHe) ........ .... .... , . .........  3.6 

Methane  (marsh-gas)  (CH<) ...  72.2 

Hydrogen  (H2) :. ..... .....  v .....  20.7 

100.4 


16 


FUELS 


THE  ARTIFICIAL  FUELS 

These  include  charcoal,  coke  and  producer  gas,  all  made  from  natural 
fuels. 

Charcoal. — Wood,  packed  in  a  kiln,  and  permitted  to  partly  burn, 
changes  into  charcoal  by  distillation  of  the  volatile  portion  by  the  heat 
produced  from  the  portion  burned.  The  charcoal  retains  the  form  of  the 
wood  from  which  it  was  made,  but  has  a  specific  gravity  of  only  0.2.  It  is 
of  a  dull-black  color,  soils  the  fingers  but  slightly  if  of  good  quality,  but 
much  if  poor.  It  should  ring  when  struck,  and  should  show  the  annual 
rings  of  the  wood  distinctly.  The  density  of  charcoal  varies  with  that  of 
the  wood  from  which  it  was  made,  dense  woods  giving  a  dense  charcoal. 
A  heaped  bushel  (1.5555  cu.  ft.)  weighs  14  to  16  Ib.  When  apparently 


FIG.  8.- — Section  of  Charcoal  Kiln. 

quite  dry,  charcoal  still  contains  10  per  cent  or  more  of  moisture.  Dry 
charcoal  contains  95  per  cent  carbon,  1.5  ash,  and  has  a  calorific  power  of 
7610  pound-calories  per  pound.  Charcoal  is  used  in  iron  blast-furnaces, 
particularly  in  localities  where  wood  is  abundant;  and  it  produces  a  pure, 
strong  iron,  free  from  sulphur,  called  "  charcoal-iron."  Charcoal  has  been 
used  also  for  silver-lead  and  copper  smelting  in  districts  difficult  of  access. 
In  these  cases  it  has  done  especially  well  when  coke  could  be  secure^  to 
use  in  conjunction  with  it.  It  is,  however,  a  friable  fuel,  making  fine  dust 
sometimes  to  the  extent  of  10  per  cent;  and  this  "  fine  "  is  apt  to  make 
trouble  in  the  blast-furnace.  If  under-burned,  it  is  heavier  and  more 
dense,  and  has  a  brown  color.  Portions  of  the  wood  found  imperfectly 
burned  are  called  "  brands  "  and  are  returned  for  the  next  burning. 

Charcoal  is  generally  made  in  a  kiln.  One  of  these  in  section  in  Fig. 
8,  shows  the  method  of  filling.  The  kiln  is  set  at  the  foot  of  a  steep  bank, 
so  that  it  can  be  charged  conveniently  from  above.  It  has  two  charge- 
doors  A  and  B.  The  first  of  the  wood  is  conveyed  through  the  lower  door, 


CHARCOAL  17 

and  placed.  The  remainder  is  brought  along  the  runway  C,  and  introduced 
through  the  upper  door  B.  There  are  three  rows  of  openings,  3  by  4  in. 
in  size,  spaced  2  ft.  apart,  around  the  bottom  of  the  kiln.  The  kiln  is 
lighted  at  the  lower  door,  and  when  fairly  started,  both  openings  A  and  B 
are  closed  with  sheet-iron  doors.  These  are  tightly  luted  with  clay,  and 
the  air  is  thus  caused  to  enter  by  the  small  holes.  When  combustion  has 
progressed  sufficiently,  these  openings  are  tightly  closed,  and  the  kiln  is 
permitted  to  cool  slowly.  The  period  of  charring  or  burning  is  eight  days 
and  the  cooling  four  days  additional.  Such  a  kiln  holds  25  cords  of  wood 
and  produces  1125  bu.  of  charcoal  weighing  16  Ib.  per  bushel,  or  about  20 
per  cent  of  the  weight  of  wood  charged. 

By-product  Charcoal. — An  example  of  the  modern  method  of  making 
by-product  charcoal  for  iron  blast-furnace  use  is  one  at  the  Pioneer  Iron 
furnace,  Marquette,  Mich.  Here  there  are  86  kilns  each  holding  8  cords. 
The  daily  requirement  is  20  carloads,  of  16  cords  each,  amounting  to  320 
cords.  The  kiln  is  packed  full  of  wood,  the  sheet-iron  doors  put  on  and 
closed,  and  fire  is  started  at  a  manhole  in  the  apex  of  the  dome.  As  soon 
as  combustion  gains  sufficient  headway,  this  opening  is  closed,  and  smoke 
escapes  by  way  of  a  flue  leading  from  the  base  of  the  kiln  to  the  chimney, 
continuing  thus  until  most  of  the  aqueous  vapor  has  escaped.  At  this 
stage  the  chimney  is  closed,  and  the  vapors  pass  by  a  smoke-main  to  the 
condensers,  the  current  being  aided  by  an  electrically  driven  fan.  The 
cold  surface  of  the  copper  tubes  of  this  condenser  precipitates  the  con- 
densible  portion  of  the  gas,  while  the  gas  itself  goes  on  to  the  boilers, 
where  it  is  burned  for  steam-making.  The  condensible  portion,  amount- 
ing to  41  per  cent  of  the  weight  of  the  wood,  is  called  green  liquor 
or  pyroligneous  acid,  and  consists  mostly  of  water,  but  contains  also 
alcohol,  tar,  ammonia  compounds,  acetone,  and  acetic  acid.  The  tar 
is  separated  in  settling  tanks,  and  the  liquor  passes  to  the  primary 
still-house.  Copper  stills  here  remove  the  vapors  of  alcohol,  acetic  acid, 
and  much  water  from  the  liquor.  The  neutralizing  tank  receives  the 
product,  and  into  this  is  mechanically  stirred  milk-of-lime  to  neutralize 
the  acid  by  the  formation  of  acetate  of  lime.  The  neutralized  liquor  is 
allowed  to  settle,  and  the  supernatant  solution  is  drawn  off  and  conveyed 
to  the  refining-still  house.  By  fractional  distillation  a  crude  wood  alcohol 
is  obtained  here,  and  a  solution  of  acetate  of  lime  is  left  behind,  and 
recovered  by  evaporating  the  solution.  The  crude  alcohol  is  then  purified 
by  further  distillation  until  a  clear  95  per  cent  wood  alcohol  is  obtained. 
A  cord  of  wood  (4500  Ib.),  yields  880  Ib.,  or  19.5  bush,  of  charcoal  of  20  Ib. 
per  bushel,  208  gal.  of  pyroligneous  acid,  8  gal.  of  wood-tar,  64  Ib.  gray 
acetate  of  lime,  and  4  gal.  wood  alcohol.  By  the  sale  of  the  wood  alcohol, 
acetate  of  lime  and  formaldehyde,  and  by  the  superior  quality  and  con- 
sequently higher  price  of  charcoal-iron,  it  has  been  possible  to  build  up 


18 


FUELS 


this  industry,  where  the  supply  of  wood  is  abundant,  in  spite  of  the  serious 
competition  of  iron  smelted  in  blast-furnaces  using  coke. 

Coke.— This  is  made  from  coal  in  kilns,  in  a  way  similar  to  that  of 
making  charcoal.  Bituminous  coal  which  cokes  or  fuses  at  the  high 
temperature  of  the  kiln  or  oven  is  used  for  this  purpose. 

The  raw  screenings,  in  the  example  below,  contained  much  fine  passing 
a  IJ-in.  bar-screen.  From  this,  the  residue  left  after  removing  the  lumps 
of  merchantable  coal,  coke  was  made.  By  "  washing,"  the  fixed  carbon 
was  increased  and  the  ash  in  the  coke  reduced  to  14.24  per  cent.  A  part 
of  the  sulphur  also  was  removed  thereby.  The  refuse  was  high  in  ash, 
and  low  in  fixed  carbon,  as  was  to  be  expected;  but  the  yield  of  washed 
coal  was  85  per  cent  of  the  raw  screenings,  and  the  coke  70  per  cent  of 
the  washed  coal.  When  the  coal  contains  slate,  "  bone,"  or  pyrite,  it  is 
improved  by  this  process  of  washing,  or  separating  the  waste-matter 
by  concentrating.  An  example  of  a  semi-bituminous  southwestern  coal  is 
shown  below: 


Moisture. 

Volatile 
Combustible 
Matter. 

Fixed 
Carbon. 

Ash. 

Sulphur. 

Raw  screenings  

1.40 

19.79 

60.25 

17.33 

0.85 

Washed  coal             .... 

0.79 

19.10 

69.35 

10.24 

0.52 

Coke  

0.43 

1.39 

83.47 

14.24 

0.82 

Refuse  or  waste  

2.22 

15.76 

30.96 

50.12 

0.93 

Composition  of  Coke. — The  ash  in  coke  varies  from  10  per  cent  to  22 
per  cent  and  the  fixed  carbon  from  77  per  cent  to  89  per  cent.  In  coke, 
high  in  ash,  not  only  has  the  ash  to  be  smelted,  but  the  fixed  carbon  is 
correspondingly  low,  so  that  such  coke  is  less  efficient.  A  great  difficulty 
with  high-ash  coke  is  that  it  is  often  friable,  making  accretions  or  scaffolds 
in  the  shaft  of  the  blast-furnace.  Analyses  of  two  typical  samples  of  bee- 
hive coke  give  the  following : 

Connellsville  Coke:  fixed  carbon,  87.5  per  cent;  ash,  11.3  per  cent; 
sulphur,  0.7  per  cent.  El  Moro  coke:  fixed  carbon,  77  per  cent;  ask,  22 
per  cent  (when  the  coke,  as  in  this  case,  is  made  from  unwashed  coal); 
sulphur,  0.9  per  cent. 

The  Coke  Ash. — In  computing  a  furnace  charge,  this  is  taken  into 
account.  Ash  of  Connellsville  coke  contains  SiO2,  44.6  per  cent;  Fe, 
15.9  per  cent;  CaO,  7  per  cent;  MgO,  1.9  per  cent.  Ash  of  El  Moro  coke 
has  SiO2,  84.5  per  cent  and  Fe  5  per  cent.  It  will  be  seen  that  this  latter  ash 
has  a  large  excess  of  silica  to  be  fluxed,  and  is  accordingly  less  desirable.  On 
the  basis  of  11.3  per  cent  ash  in  the  Connellsville  coke  we  would  have  SiC>2, 
5.0  per  cent;  Fe,  1.8  per  cent;  CaO  and  MgO,  1  per  cent  of  its  total  weight. 


COKE 


19 


Beehive  Coke. — A  beehive-oven  (see  Fig.  9  at  B),  is  charged  through  a 
hole  in  the  roof.  Each  oven  holds  5  to  6  short  tons  of  coal.  In  Pennsyl- 
vania an  oven  yields  per  week  two  charges  of  48-hour  coke  and  one  of  72- 
hour.  The  charge  in  making  72-hour  coke  is  dropped  in  the  morning  into 
the  hot  oven  from  a  coal  larry  or  car  above,  and  is  leveled  through  the  side 
door,  filling  the  oven  to  the  depth  of  26  in.  The  door  is  then  walled 
up  with  dry  brick  and  plastered  over,  but  an  opening  is  left  near  the  top, 
as  shown  in  section,  for  the  admission  of  air.  Combustion  from  the  red- 
hot  brickwork  soon  begins,  and  a  dark  smoke  escapes  at  the  top  opening. 
After  four  hours  this  becomes  dense  and  white,  and  the  gases  ignite  or 
strike,  and  flames  issue  from  the  top.  For  twelve  hours  the  oven  burns 
with  a  dull,  smoky  flame  above  the  surface  of  the  charge.  The  flame  be- 


n 

n 

L 

\/q 

I-" 

:    ^q 

r 

-  -  i 

E 

•  t  T 

E 

•  -.'.£  n 

; 

__  — 

c 

I  \ 

1 

H  m. 

A 

FIG.  9. — Sections  of  By-product  (A)  and  Bee-hive  Coke  Ovens. 

comes  bright  by  the  second  day  and  then  the  air-supply  is  partly  cut  off. 
On  the  third  day  still  less  air  is  admitted,  and  at  the  end  of  this  day  no 
more  flames  appear  and  the  whole  interior  of  the  oven  is  red-hot.  The 
air-openings  are  now  luted,  and  the  charge  is  left  in  this  condition  until 
the  morning  of  the  fourth  day  when  the  coke  is  drawn.  The  actual  coking 
is  complete  in  fifty-five  hours,  and  the  whole  operation,  from  one  charging 
to  the  next,  in  seventy-two  hours.  To  draw  the  coke  the  temporary  brick 
wall  of  the  door  is  taken  down,  and  water  from  a  hose  played  into  the  oven. 
After  being  thus  cooled  on  the  surface,  the  coke  is  pulled  out  with  a  long- 
handled  coke-drag  or  hook,  and  further  cooled  with  water  while  being 
withdrawn. 

The  process  of  fusing  and  coking  begins  at  the  top,  and  extends  down- 
ward through  the  mass  of  coal  to  the  bottom  of  the  oven,  and  the  coke, 
when  well  burned,  takes  the  form  of  prismatic  masses,  see  Fig.  9,  with 


20  FUELS 

hard  side-surfaces  of  a  silvery  steel-gray  color,  and  top  ends  soft  and 
nearly  black.  The  silvery  appearance  is  due  to  deposited  carbon,  which 
has  the  desirable  quality  of  protecting  the  coke  against  the  action  of  the 
furnace-gases.  The  black  ends,  on  the  contrary,  are  readily  attacked. 
A  good  coke  has  a  well-developed  cell  structure  which  permits  the  pene- 
tration of  the  hot  ascending  gases  in  the  blast-furnace.  This  so  raises  the 
temperature  of  the  coke  that  the  air,  at  the  bottom  of  a  furnace  striking  it, 
produces  vigorous  and  rapid  combustion.  Other  qualities  are  purity, 
uniform  quality,  and  sufficient  coherence  for  handling.  Purity  depends 
upon  a  low  ash,  10  per  cent  being  good,  and  6  or  8  exceptionally  pure. 
Coke  intended  for  iron  blast-furnace  work  should  not  contain  more  than  1 
per  cent  sulphur  and  commonly  less  than  0.5  to  0.8  per  cent  of  this  element. 
For  lead  or  copper  blast-furnaces  high  sulphur  does  not  greatly  matter. 
"  Uniform  quality  "  means  but  a  small  amount  of  "  black  ends."  These 
as  stated,  burn  in  the  upper  part  of  the  iron  blast-furnace  by  the  action  of 
carbon-monoxide  gas.  "  Coherence  in  handling  "  as  is  evident,  is  impor- 
tant where  coke  must  be  transported  far,  and  rehandled  at  the  smelting 
works.  Fines  tend  to  "  slow  down  "  a  blast-furnace,  but  can  be  rejected  by 
the  use  of  a  coke-fork.  The  calorific  value  of  Pittsburg  coke,  containing 
89  per  cent  fixed-carbon,  10  per  cent  ash,  and  1  per  cent  sulphur,  is  7272 
Ib.-cal.  per  pound. 

The  By-product  Coke  Oven. — This  is  coming  increasingly  into  use, 
due  to  the  fact  that  in  it,  the  products,  aside  from  the  coke,  can  be  saved 
and  sold  to  advantage,  and  not  wasted  as  is  done  in  making  beehive  coke. 
Moreover,  due  to  the  constant  high  and  quick  heat  produced,  it  can  coke  a 
coal  that  contains  but  little  fusible  matter  or  is  nearly  non-coking. 

.A  beehive  oven  will  yield  2000  Ib.  of  coke  from  3200  Ib.  of  coal,  a 
by-product  oven  from  the  same  amount  of  coal  will  yield  2300  Ib.  coke; 
19  gal.  tar;  42  Ib.  ammonium  sulphate;  4.5  gal.  benzol  (motor  fuel)  and 
10,000  cu.  ft.  of  fuel  gas. 

The  types  of  ovens  of  this  kind  extensively  in  use  are  the  Otto-Hoffman, 
the  Semet-Solvay,  and  the  Koppers.  The  longitudinal  section  of  a 
Semet-Solvay  oven  is  shown  in  Fig.  10.  The  cross-section  through  the 
ovens  themselves  is  given  in  Fig.  9.  i 

Coal  is  brought  in  over  two  tracks,  and  discharged  into  feed-hoppers. 
It  is  drawn  from  these  as  required,  and  conveyed  to  two  sets  of  rolls,  one 
for  coarse,  the  other  for  fine  crushing,  and  reduced  to  a  size  of  4  to  10  rnesh. 
The  crushed  coal  is  raised  by  an  inclined  elevator,  and  discharged  into  the 
rr-^in  storage  coal-bin.  This  bin  has  a  hopper-shaped  bottom  with  several 
discharge  spouts,  delivering  to  an  8-ton  larry  which  runs  along  on  top  of 
the  ovens  or  retorts,  of  which  there  may  be  20  to  60,  placed  side  by  side,  in 
one  block  of  masonry.  Each  retort  or  coking  chamber  is  17  in.  wide,  43 
ft.  6  in.  long,  and  6  ft.  6  in.  high,  and  is  closed  at  each  end  by  an  air-tight 


BY-PRODUCT  OVENS 


21 


cast-iron  door.  In  Fig.  9  is  shown  a  transverse  section  of  such  a 
chamber  with  the  interesting  lines  of  fractures  and  columnar  structure  of 
the  coke  indicated. 

At  Fig.  10,  the  charge-car  is  seen  above  the  retorts.  It  is  worked  by  an 
electric  motor  and  consists  of  4  hoppers  supported  by  a  frame  upon  a 
traveling  carriage.  The  doors  of  the  chamber  being  closed,  and  the 
chamber  itself  hot  from  previous  operation,  a  charge  of  8  tons  of  coal  is 
dropped  in,  and  leveled  by  means  of  the  top  bar  of  the  charging  machine, 
Fig.  12,  inserted  through  an  opening  near  the  top  of  the  door.  Distillation 


Pusher 

H-.i'l 


FIG.  10. — Semet-Solvay  By-product  Oven. 

at  once  begins,  and  the  gases  are  conducted  to  condensing-chambers  to 
free  them  from  certain  by-products,  such  as  tar,  ammonia,  and  benzol. 
The  first  portion  of  the  gas  is  highest  in  illuminating  power,  say  24  candle- 
power,  but  later  drops  to  16  candle-power.  The  first  is,  therefore,  sent  to 
the  city-mains  for  use  as  illuminating  gas,  the  latter  reserved  to  heat  the 
chambers  by  combustion  in  flues  which  encircle  them.  These  flues  are 
beneath  the  chambers,  and  the  side-walls  are  constructed  to  provide  them 
for  heating  the  oven,  and  maintaining  the  activity  of  the  distillation.  The 
products  of  combustion,  before  entering  the  stack,  go  through  a  regenerating 
chamber  containing  a  checker-work  of  tile,  while  air  is  preheated  for  com- 
bustion in  a  similar  chamber  at  the  other  side.  Thus  the  gas  is  burned 
with  highly  heated  air,  and  produces  an  intense  heat  in  the  walls  of  the 
coking-chambers.  The  reversing  valves  are  now  changed,  and  the 
currents  of  air  and  gas  caused  to  move  in  the  opposite  direction.  The 


22 


FUELS 


FIG.  11. — Koppers  By-product  Coke-oven. 


FIG.  12. — Perspective  View,  Ovens,  Coke-pusher  and  Leveler. 


GAS  PRODUCERS 


23 


direction  is  thus  repeatedly  alternated,  as  is  customary  in  open-hearth 
work.  At  the  end  of  twenty-four  hours,  when  coking  is  complete,  the  end 
doors  are  opened  and  the  coke  is  pushed  out  by  means  of  a  coke  pusher, 
Fig.  12.  The  pusher-head  is  shown  at  the  left  in  Fig.  10.  The  coke  is  received 
in  coke  car,  shown  at  the  right  of  the  oven,  and  is  here  cooled  with  water. 
The  total  yield  of  coke  is  72  per  cent,  or  6  per  cent  more,  for  the  same 
coal,  than  that  of  a  beehive  oven.  The  coke  is  hard,  dense,  and  as  reliable 
as  beehive  coke  made  from  the  same  coal,  but  has  not  the  silvery  gloss  of 
the  latter. 

Costs. — The  actual  cost  of  making  coke  may  be  stated  as  50  cents  per 
ton  in  the  beehive  process  and  37  cents  in  by-product  ovens.  To  this 
must  be  added  the  cost  of  the  1J  tons  of  coal  required.  A  beehive  plant 
operated  six  days  per  week  and  of  400-ton  daily  capacity  would  cost 
$80,000.  A  by-product  plant  of  the  same  capacity  would  cost  $300,000. 
Allowing  for  interest  and  depreciation,  the  cost  is  found  to  be  much  the 
same  for  either  process. 

PRODUCER-GAS 

Of  the  various  kinds  of  producers  used  for  making  artificial  fuel-gas  we 
shall  consider  two,  "  the  simple  producer  "  and  the  "  mixed-gas  producer." 


FIG.  13. — Section  View  of  Gas-producer  (hand  poked). 

The  Simple  Producer. — These  use  ordinary  or  inferior  fuels,  such  as 
ttood,  wood-refuse,  bark,  sawdust,  or  peat,  but  generally  soft  or  hard 
X>al.  We  have  shown  in  Figs.  176  and  177,  in  the  sections  of  furnaces 


24  FUELS 

containing  fuel,  how  gas  is  produced  where  air  rises  through  a  deep  coke 
fire  and  where  fuel  is  thus  in  excess. 

Fig.  13  is  a  simple  producer,  the  necessary  air  being  supplied  by  a 
natural  draft  or  by  a  fan.  The  fuel,  descending  in  the  producer,  first  is 
dried  by  the  hot,  rising  gases,  then  further  heated  until  the  volatile  matter 
is  distilled,  and  finally,  as  it  reaches  the  lowest  zone,  is  oxidized  or  burned 
by  the  entering  air.  The  residue  is  the  ash  of  the  fuel,  which  is  withdrawn 
at  the  bottom.  The  escaping  gases  issue  at  a  temperature  of  300°  to 
1000°  C. 

In  the  operation  of  a  hand-poked  producer,  the  coal,  when  charged, 
if  left  to  itself  would  soon  burn,  leaving  holes  in  the  fuel-bed,  through 
which  would  come  up  unconsumed  air.  To  avoid  this  a  long  bar  is  run 
down  through  one  of  the  poke-holes  shown  in  the  producer  top  to  break 
up  the  hung-up  coal  and  again  make  the  bed  continuous. 

An  analysis  of  producer-gas  made  from  soft  coal  gave  the  following 
results  by  volume: 

Per  Cent. 

Carbon  dioxide  (CO2) 5.0 

Carbon  monoxide  (CO) 23 . 0 

Oxygen  (O2) 0.5 

Ethylene  (C2H4) 0.5 

Methane  (CH4) 3.0 

Hydrogen  (H2) 10.0 

Nitrogen  (N2) 58.0 


100.0 

Each  pound  of  coal  will  give  60  cu.  ft.  of  such  gas,  having  a  heating  value 
of  82  Cal.  per  cubic  foot. 

The  Mixed-gas  Producer.  —  This  is  the  producer  commonly  used.  In 
it  some  steam  or  water  vapor  is  blown  with  the  air  into  the  burning  fuel, 
and  there  reacts  upon  the  carbon  as  follows  : 


58,000     29,000  =-29,000 

One  volume  of  steam  makes  one  volume  of  carbon  monoxide  and  one  of 
hydrogen.  The  steam  may  be  obtained  from  the  water-soaked  ashes  by 
evaporation  in  the  lower  part  of  the  producer,  or  as  in  Fig.  13  may  be 
injected  under  pressure  into  the  fire.  Steam  also  disintegrates  the  clinker 
and  facilitates  its  removal.  The  carbon  dioxide  formed  means  the  pro- 
duction of  heat  later  absorbed  in  the  formation  of  water-gas  and  hence  is 
;ji  a  way  useful,  since  the  formation  of  water-gas  can  be  carried  farther. 

The   Hughes   Mechanically  Poked   Continuous   Gas-producer.  —  Fig. 
14  is  a  sectional  elevation  of  a  plant  containing  a  row  of  Hughes  producers 


GAS  PRODUCERS 


25 


which  give  a  mixed  gas  to  supply  open-hearth  furnaces  used  in  making 
steel. 

The  special  feature  of  this  type  of  producer  is  the  mechanical  poker 
F,  which  is  a  water-cooled  steel  casting  suspended  and  secured  to  a  shaft  S. 
The  poker  is  actuated  by  a  mechanism  which  moves  the  poker  back  and 
forth,  agitating  and  breaking  up  the  mass  of  fuel,  in  the  slowly  rotating 
shell,  evenly  distributing  the  coal,  and  helping  to  work  the  ashes  down- 


FIG.  14. — Hughes  Mechanically  Poked  Continuous  Gas-producer  Plant. 

ward.  Thus  the  labor  of  hand-poking  is  eliminated  and  the  fuel  is  regu- 
larly stirred.  This  uniform  treatment  has  proved  of  great  advantage, 
giving  uniform  results  in  quality,  quantity,  and  supply  of  gas,  with  a 
reduction  of  operating  costs. 

The  brick-lined  producer  shell  G,  or  body,  is  of  steel  having  a  cast- 
iron  base-ring  and  a  cast-iron  water-sealed  ash-pan  J  and  a  turn-table  R, 
ill  bolted  together  in  one.  The  turn-table  R  has  at  its  outer  circumfer- 
ence a  cast-iron  rack  into  which  meshes  a  spur  pinion  keyed  to  the  vertical 


26 


FUELS 


shaft  S  and  connected  to  the  horizontal  main  shaft  by  a  train  of  gearing. 
Thus  the  turn-table  is  rotated  and  with  it  the  ash-pan,  base-ring,  and  body 
G.  The  bottom  of  the  turn-table  is  fitted  with  a  steel  tread  resting  on  six 
conical  chilled-iron  carrying-wheels.  As  the  producer  slowly  revolves,  the 


FIG.  15. — Cross-section  of  Revolving  Eccentric  Gas-producer. 

ashes  work  down  into  the  water-sealed  ash-pan  from  which  they  are 
shoveled  directly  into  a  car  in  order  to  remove  them.  The  steel  producer 
top  is  secured  rigidly  to  the  floor  structure  and  a  water  seal  is  formed  by 
a  flange  at  the  outer  circumference  of  the  producer  cover.  Ifc  carries  the 
poker  mechanism,  two  charge-hoppers  P,  and  the  gas  outlet  or  off-take,  T. 


GAS  PRODUCERS 


27 


The  latter  leads  to  a  gas  main  or  flue  connected  to  the  whole  row  of  pro- 
ducers. The  base  of  the  producer  has  a  blast  inlet-pipe  H  with  a  cast- 
iron  deflecting  plate  for  covering  the  air  opening.  Air  is  delivered  to  this 
pipe  by  means  of  a  blower.  The  producer  is  generally  driven  by  an 
electric  motor  requiring  three  electric  horse-power. 

A  hand-poked  producer  has  a  capacity  of  10  Ib.  of  coal  per  square  foot 
of  grate  area  per  hour,  while  the  mechanically  poked  producer  can  burn 
on  an  average  25  Ib.  This  figures  out  approximately  one  ton  of  coal  per 
hour  for  the  Hughes  producer  of  10  ft.  internal  diameter.  An  average 
quality  of  gas  can  be  maintained  of  a  composition  as  follows: 


CO2  CO 

4  per  cent     25  per  cent 


Hydrocarbons 
3  to  4  per  cent 


H 

13  per  cent 


N 
53  per  cent 


FIG.  16. — Loomis-Pettibone  Gas-making  Plant. 


Revolving  Eccentric  Gas  Producer.--Fig.  15  illustrates  a  plant 
containing  a  double  row  of  gas  producers  having  grates  eccentrically  set 
so  that  as  they  revolve  they  carry  the  fuel  reciprocally  to  and  from  the 
interior  walls  of  the  producer  shell.  The  hoppers  are  filled,  and  the 
charge  dropped  as  in  the  bell  of  an  iron  blast-furnace.  By  the  revolution 
of  the  grate  the  ashes  work  to  and  under  the  peripheral  edge  of  the  pro- 
ducer shell.  Formerly  needing  1.75  tons  of  coal  to  give  gas  enough  to 
smelt  1  ton  of  charge  in  the  open-hearth,  the  producer  can  now  do  the 
same  work  with  a  consumption  of  but  0.7  to  0.9  ton. 

The  Loomis-Pettibone  Gas  Apparatus. — Fig.  16  shows  a  complete 
plant  of  the  Loomis-Pettibone  system,  with  a  positive  gas  exhauster.  It 
is  intended  both  for  producer-  and  water-gas.  Its  operation  is  as  follows: 
Hot  fires  are  burning  in  both  producers  or  generators,  and  the  gas  exhauster 


28  FUELS 

is  in  operation.  Air  is  now  drawn  upward  through  generator  1,  burning 
the  fuel  and  making  producer-gas.  This  generator  may  have  just  received 
fresh  coal  at  E,  and  the  coal-smoke,  tarry  matter,  and  producer-gas  from 
it,  are  together  drawn  down  through  the  hot  fire  in  generator  2,  being  com- 
pletely burned  and  fixed 'in  so  doing.  The  gas  now  goes  through  valve  B 
to  the  boiler  (valve  A  being  closed),  and  the  heat  is  there  absorbed.  11 
then  passes  from  the  top  of  the  boiler  through  the  pipe  shown  to  the  bottom 
of  the  "  scrubber/'  a  cylindrical  tower  of  sheet-steel,  in  which  it  is  caused 
to  pass  upward  through  pieces  of  coke  resting  upon  perforated  trays 
The  coke  here  is  kept  wet  by  means  of  a  water-spray,  and  the  gas  is  thereby 
cooled  and  cleaned.  The  water  drains  off  by  the  water-sealed  pipe  V. 
Rising  to  the  top  and  to  the  wider  part  of  the  tower,  the  gas  passes  through 
a  layer  of  fine  shavings  or  "  excelsior,"  to  remove  any  remaining  dust.  11 
is  then  drawn  through  the  Root  positive-blast  exhauster  W,  and  finally  is 
driven  through  pipe  Z  to  the  gasometer  for  producer-gas,  where  it  is  storec 
for  use.  The  fire  in  generator  1  having  become  clear  and  hot,  generator  e± 
is  charged  afresh,  and  the  ash-pit  door  opened.  The  gas  current  is  ther 
charged  from  generator  2  to  generator  1,  through  valve  A  (valve  B  having 
been  shut)  to  the  boiler,  thence  through  the  scrubber  and  exhauster  W 
to  the  gasometer.  The  direction  of  the  current  is  thus  changed  at  inter 
vals.  For  making  water-gas,  the  ash-pit  door  is  closed  and  steam  fron 
the  boiler  is  injected  beneath  the  grate  of  the  generator  while  the  fire  i; 
hot.  The  formation  of  the  water-gas  is  completed,  or  the  gas  is  "  fixed  ' 
by  causing  it  to  pass  down  through  the  other  generator,  it  having  beer 
found  that  a  part  of  the  hydrogen  reverts  to  steam  without  so  doing 
The  making  of  water-gas  cools  the  fire  and  after  a  few  minutes  the  stean 
must  be  shut  off  and  air  again  substituted.  While  water-gas  is  being  made 
it  may  go  to  the  gasometer  through  the  pipe  Z,  or,  if  desired,  to  permit  I 
to  go  to  the  water-gas  holder,  Z  may  be  closed  and  Y  opened.  When  kep 
separate,  water-gas  is  reserved  for  certain  heating  operations  for  whicl 
producer-gas,  of  lower  calorific  power,  would  be  unsuited.  The  purge 
pipe  X  is  opened  when  starting,  and  by  this  means  air  in  the  system  i 
expelled  before  gas  is  turned  into  the  gasometer.  Steam  may  also  b< 
admitted  above  the  fire,  and  thus  caused  to  pass  down  through  the  generate 
and  form  water-gas.  In  fact,  both  air  and  steam  may  be  introduced,  eithe 
below  or  above  the  fires,  to  suit  the  best  conditions  of  operating. 

Comparing  the  two  systems,  the  hand-poked  producer  costs  70  cents 
while  the  Hughes  producer  can  be  operated  for  50  cents  per  ton  of  coa 
burned.  At  the  same  time  a  Hughes  mechanically  poked  producer  instal 
lation  is  estimated  to  cost  $38,960  as  against  $45,200  for  a  hand-operatec 
one. 


PULVERIZED  COAL 


29 


PULVERIZED    COAL 

This  is  prepared  from  run-of-mine  or  from  slack  coal.  It  is  delivered 
by  car  into  storage  bins,  whence  it  is  drawn  off  upon  a  conveying  belt  to 
the  feed  hopper  of  a  dryer,  Fig.  72.  The  head  pulley  at  the  delivery 
end  of  the  belt  is  magnetized,  so  that  stray  pieces  of  iron  and  steel  are 
removed.  In  this  way  trouble  is  avoided  in  the  grinding  machines  later  on. 

D^ied  down  to  1  per  cent  moisture  the  dryer  discharge  feeds  to  a  slightly 
corrugated  roll  set  to  reduce  it  to  pea  size.  As  fast  as  crushed  this  product 
is  raised  by  a  belt  elevator  to  a  stock-bin  to  be  drawn  off  as  desired 
to  the  Raymond  roller  mill,  where  it  is  to  be  pulverized. 


FIG.  17. — Raymond  Roller  Mill 
(belt  driven). 


FIG.  18. — Raymond  Roller  Mill. 


Fig.  17  is  a  view  of  the  roller  mill,  where  at  the  left  is  the  spout  from 
the  stock-bin,  having  at  its  foot  a  deeply  corrugated  feed  roller  called 
"  star  feed."  This,  as  it  revolves,  gives  a  regulated  supply  within  the 
truncated  casing,  where  the  grinding  is  performed.  As  in  the  Huntington 
mill,  there  is  a  "  bull  ring,"  between  which  and  the  three  suspended 
rolls  the  grinding  is  performed.  A  head  on  the  central  vertical  shaft  has 
three  suspended  shafts  carrying  the  rolls,  so  that  in  rapid  rotation  they  are 
strongly  pressed  against  the  inside  of  the  bull  ring,  quickly  pulverizing  the 
coal.  The  suction  pan  above  draws  up  the  pulverized  material,  delivering  it 
tangentially  to  the  large  collecting  cone  at  the  right,  where  it  is  whirled  in 
cyclone  fashion  to  the  periphery  to  settle  to  the  point  of  the  cone.  The 
air,  thus  freed  from  the  bulk  of  its  contained  dust,  is  returned  by  the  pipe 
rising  from  the  top  of  the  cone  to  the  lower  large  exterior  casing,  and  passes 


30  FUELS 

upward  inside  the  bull  ring.  The  coarse  particles,  that  have  escaped 
grinding,  and  fallen  to  the  bottom  are  lifted  by  plows  that  throw  them  up, 
and  with  the  aid  of  the  upward  wind  current,  again  lift  them  to  the  grind- 
ing zone. 

The  pulverized  coal  from  the  roller  mill  is  taken  by  a  screw  conveyer 
to  a  pulverized  coal-bin  of  25  tons  capacity.  Thence  it  goes  by  other 
conveyors  to  the  respective  roasters.  The  coal  must  be  so  finely  powdered 
that  85  per  cent  of  it  is  of  minus  200  mesh,  since  the  finer  it  is  ground, 
the  greater  its  efficiency. 

The  method  of  feeding  powdered  coal  to  a  reverberatory  furnace  is 
described  under  head  of  "  pulverized  coal-firing." 

The  Holback  Powdered  Coal  Distributing  System. — This  comprises 
an  air-supply  pipe  into  which,  under  fan  pressure  powdered  coal  is  fed 
by  feed-screws  from  coal-storage  bins  branching  to  supply  all  the  furnaces. 
A  return  pipe  takes  the  excess  of  air,  together  with  the  unused  coal  back  to 
the  storage  bins.  As  more  air  is  used  so  the  coal  supply  is  automatically 
increased  at  the  feed  screws  in  direct  proportion.  A  control  valve  is 
provided  at  each  branch  as  also  a  burner  for  the  coal  and  air  to  each 
furnace. 


CHAPTER  III 

REFRACTORIES 

REFRACTORY  MATERIALS  AND  THEIR  PROPERTIES 

General. — The  foundations  of  a  furnace  may  be  of  concrete  or  of  stone 
aid  in  lime-mortar,  the  moderately  heated  exterior  of  common  building 
brick  also  laid  in  lime-mortar,  but  for  the  interior  lining  it  is  necessary  to 
use  refractory  material  to  withstand  the  high  temperature  and  to  resist 
the  scouring  and  corroding  action  of  the  molten  contents  of  the  furnace. 
At  a  temperature  below  a  red  heat  the  combined  moisture  of  lime-mortar 
would  be  expelled,  and  the  mortar  in  consequence  would  crumble.  At  a 
dull  red  heat  many  stones  crack  and  flake  off  at  the  surface  of  irregular 
expansion.  Sandstone,  however,  is  resistant  to  fire,  and  has  been  used  for 
furnace  lining.  Red  bricks,  laid  in  clay  mortar,  withstand  a  moderate 
red  heat,  but,  at  a  temperature  much  above  this,  begin  to  soften  or  melt. 

Refractories. — These  substances  are  infusible  at  the  high  temperatures 
for  which  they  are  intended.  Thus  firebrick  only  begins  to  soften  at  1500° 
to  1600°  C.,  and  silica  brick  at  1600°  to  1700°  C.  Refractories  may  be 
divided  into  the  three  following  classes : 

Acid  (Silica-brick,  Sand  and  Canister). — These  are  used  to  resist  the 
scouring  or  corrosive  action  of  acid  slags.  Being  highly  refractory  they 
are  more  generally  used  for  roofs  or  arches  exposed  to  the  highest  tem- 
peratures. In  such  positions  out  of  contact  with  the  molten  contents  of 
furnaces  they  are  not  required  to  resist  a  serious  fluxing  action. 

(2)  Neutral  (Graphite,  Chrome-iron,  Fireclay,  Bone-ash,  and  Carbon- 
brick). — These  materials  well  resist  the  action  of  neutral  slags  which  are 
neither  basic  nor  acid.     In  the  case  of  a  basic  open-hearth  furnace,  for 
example,  it  is  customary  to  interpose  a  layer  of  neutral  chrome-iron  brick 
between  the  roof  cf  silica-brick  and  the  basic-lined  hearth  slightly  above 
the  level  of  the  surface  of  the  molten  contents  of  the  furnace  where  it  would 
be  unaffected  by  it.     Were  silica-brick   used  in  contact  with  the  basic 
lining,  they  would  react  upon  the  lining  and  melt. 

(3)  Basic  (Dolomite,  Magnesite,  etc.) — These  are  used  where  the  slag 
or  matte  is  basic,  as  in  the  hearth  of  the  basic  open-hearth  furnace.     Basic 
slags  quickly  scour  or  corrode  an  acid,  or  even  a  neutral  lining.     It  will  be 
noticed  that  all  the  foregoing  refractories  not  only  have  special  resistant 

31 


32  REFRACTORIES 

power  but  are  infusible.     This  is  particularly  the  case  with  carbon,  either 
in  the  form  of  gas-carbon  or  charcoal. 

ACID    REFRACTORIES 

Sand. — This  is  used  in  repairing  or  fettling  the  interior  borders  or  walls 
of  reverberatory  furnaces.  It  is  made  to  form  a  steep  bank  extending 
above  the  level  of  the  molten  bath,  to  protect  the  wall  from  the  corrosive 
action  of  the  molten  slag.  Repairs  are  made  after  the  charge  has  been 
withdrawn,  when  the  interior  sides  of  the  furnace  are  exposed.  In  copper 
reverberatory  work  the  sand  is  thrown  in  by  means  of  shovels,  or  placed 
by  paddles  or  spoons  provided  with  16-ft.  handles  to  allow  the  sand  to  be 
dropped  at  the  exact  spot  required.  Sometimes  a  little  clay  is  incor- 
porated with  the  sand  that  it  may  be  formed  into  balls.  These  are  skill- 
fully thrown  across  the  furnace  through  a  door  to  an  eroded  spot,  or  inserted 
by  means  of  the  paddle,  mentioned  above,  and  pressed  into  position  with 
the  bowl  of  a  long-handled  ladle.  The  bottoms  of  reverberatory  furnaces 
are  frequently  made  of  sand  in  layers,  and  each  layer  fired  upon  and 
melted  successively,  at  the  highest  temperature  of  the  furnace.  The 
sand,  fritted  together,  and  hardened  into  a  coherent  bed  in  this  way,  is 
built  to  the  thickness  of  perhaps  2  ft. 

Ganister. — This  is  used  for  furnace-  or  converter-lining  in  copper  work. 
It  is  composed  of  a  mixture  of  crushed  silicious  rock  or  quartz  to  which 
has  been  added  about  15  per  cent  clayey  material  to  make  it  cohere. 
For  acid-lined  copper  converters,  a  silicious  ore  carrying  gold  and  silver 
may  be  used  instead  of  barren  quartz  rock.  The  material  is  rapidly  eaten 
or  scoured  away  by  the  action  of  the  molten  charge,  and  the  precious 
metal  contained  enters  the  charge.  This  in  reality  results  in  a  kind  of 
ore-smelting,  performed  incidentally,  and  without  additional  cost. 

Silica  Brick. — When  quartz  or  sandstone,  containing  98  per  cent  silica, 
is  moistened  and  mixed  in  a  wet  pan  (Fig.  170)  with  a  little  lime  paste 
made  from  quick-lime,  it  coheres  sufficiently  to  be  molded  into  brick. 
These  are  first  dried  in  a  steam-heated  drying-room,  then  carefully  placed 
in  kilns  in  open  order,  and  burned  at  a  temperature  gradually  increasing 
to  a  white  heat.  Fig.  19  represents  a  kiln  of  the  down-draft  type.  It  is  a 
dome-shaped  oven,  18  to  30  ft.  diameter,  coal-fired  by  means  of  fireplaces 
set  in  the  exterior  wall.  The  flues  within  this  wall  are  arranged  as  shown, 
so  that  the  entering  flame  rises  to  the  crown  of  the  arch,  and,  passing  down- 
ward through  the  brick,  goes  to  the  adjoining  stack  through  flues  in  the 
floor  of  the  kiln.  Thus  a  high  even  temperature  is  obtained,  and  the  brick 
becomes  sufficiently  sintered  to  stand  handling  and  transportation,  though 
never  as  strong  as  the  fireclay  brick. 

Besides  the  lime-bond  brick,  above  described,  made  by  the  addition  of 


BRICK   MAKING 


33 


lime  to  silica,  a  clay-bond  brick,  less  refractory,  is  made  by  the  admixture 
of  four  parts  of  flint  with  one  of  clay.  This  makes  a  stronger  brick  than 
the  lime-bond.  The  composition  of  each  of  these  kinds  of  brick  is  as 
follows: 


Lime-bond 
Brick, 
Per  Cent. 

Clay-bond 
Brick, 
Per  Cent. 

SiOj           

93.48 

86   32 

AljOj                        ....•;  

3  82 

11  24 

Total  fluxing  bases 

2  62 

2  50 

99.92 

100.06 

FIG.  19.— Brick  Kiln. 

The  clay-bond  brick  shows  its  greater  fusibility  in  its  alumina  and 
silica  ratio,  as  will  be  seen  under  the  constitution  of  firebrick,  and  the 
proportion  of  alkali  is  higher  than  in  the  lime-bond  brick,  causing  it  to 
be  much  less  refractory  Silica  brick  withstands  the  highest  tempera- 
tures, and  expands  when  heated  To  provide  for  this,  expansion  joints  are 
arranged  in  the  roof,  side  walls,  and  bridge  of  reverberatory  furnaces,  which 
close  as  the  temperature  rises.  To  slack  off  the  tie-rods,  also,  is  another 
way  to  accomplish  the  same  purpose.  Without  this,  furnace  arches  would 
bulge,  and  tie-rods  would  break.  The  linear  expansion  of  these  bricks 
when  elevated  in  temperature  to  a  white  heat  is  2.5  per  cent. 


34    .  REFRACTORIES 

NEUTRAL    AND   BASIC    REFRACTORIES 

Graphite  or  Plumbago. — Pure  carbon  in  the  absence  of  air  is  permanent 
and  infusible  at  the  highest  temperatures.  This  is  well  exemplified  in  the 
carbon  filament  of  an  incandescent  lamp.  Even  in  the  arc-light,  the 
carbons,  though  gradually  consumed,  do  not  melt.  In  blast-furnaces, 
pulverous  carbon  accumulates  and  forms  scaffolds,  and  carbon-brick, 
made  of  gas-carbon,  has  been  used  with  some  degree  of  success  for  the 
bosh-lining  of  iron  blast-furnaces.  Graphite  is  essentially  carbon,  but 
contains  as  impurities  a  little  iron  and  a  small  quantity  of  gangue  sub- 
stance. An  analysis  of  Canadian  graphite  gives  2  per  cent  volatile  matter, 
20  per  cent  ash,  and  80  per  cent  carbon.  Such  graphite  is  used  for  graphite 
or  plumbago  crucibles  and  retorts,  when  mixed  with  45  per  cent  air-dried 
clay  and  5  per  cent  sand.  Graphite  in  these  mixtures  is  not  only  refractory, 
but  prevents  shrinking  and  cracking  when  the  crucible  or  other  object  is 
dried  after  being  formed. 

Chromite  or  Chrome-iron. — This  is  a  double  oxide  of  iron  and  chro- 
mium (FeOC^Og)  generally  containing  a  little  gangue.  Chrome  ore  is 
made  into  bricks  by  crushing  the  ore,  mixing  with  lime  as  in  making  silica 
brick,  and  burning.  These  bricks  should  not  contain  more  than  40  per 
cent  Cr20s.  Chromite  is  not  attacked  by  silicious  slags,  and  resists  high 
temperatures. 

Fireclay,  Firebrick,  and  Tile. — These  refractories  are  the  best  known 
and  the  most  used.  The  term  fireclay  applies  to  kinds  of  clay  capable  of 
withstanding  a  high  degree  of  heat.,  In  good  fireclay  the  total  percentage 
of  fluxing  impurities,  such  as  ferric  oxide,  lime,  magnesia,  and  the  alkalis, 
is  small  (3.5  per  cent  or  less).  In  all  fireclay  the  water  and  some  of  the 
silica  is  combined  chemically  with  the  alumina.  This  forms  a  hydrous 
aluminum  silicate,  called  kaolinite.  Further  silica  present  is  in  the  form 
of  quartz  sand.  Either  kaolinite,  or  quartz  alone,  has  a  high  fusion  point 
(1850°  C.),  but  in  mixture,  the  fusion  point  is  lower,  and  this  reaches  a 
minimum  at  1670°  when  10  per  cent  kaolinite  is  present.  By  the  con- 
tinued addition  of  silica  to  kaolinite  we  therefore  get  a  diminution  of 
refractoriness  until  this  exact  proportion  is  reached,  and  after  this,  by 
continued  addition  of  sand,  an  increase.  The  fireclay,  accordingly,  is 
most  refractory  that  contains  the  lowest  percentage  of  fluxing  base,  and 
the  least  uncombined  sand.  A  factor  further  affecting  the  refractoriness 
is  the  coarseness  of  grain.  The  New  Jersey  air-dried  clays  have  the  fol- 
lowing composition  and  refractory  qualities: 

Per  Cent.       Per  Cent. 

Kaolinite  (clay  base) 57 . 47  98 . 95 

Free  silica 40 . 09  0 . 24 

Total  fluxing  bases 2 . 53  0. 99 

100.09  100.18 


BRICK  MAKING  35 

Per  Cent.  Per  Cent. 

Si02 67.26          45.76 

A12O3 23.36          39.05 

H2O  (combined) 6.94  14.46 

(Fe2O3,  CaO,  alkalis) 2.53  0.99 


100.09         100.26 
Temperature  of  fusion 1670°  C.      1810°  C. 

The  clay  base  is  computed  as  A^Os,  2SiC>2  with  combined  water.  The 
silica  not  present  in  this  combined  form  is  regarded  as  "  free."  It  is  seen 
that  the  less  refractory  clay  (IV)  contains  more  fluxing  base,  more  silica 
and  less  alumina  than  (V)  to  account  for  its  fusibility.  The  first  (IV),  is 
harder  than  (V)  because  more  fusible,  and  is  an  acid  brick,  whereas  (V)  is 
neutral.  The  second  (IV),  is  a  type  of  most  of  the  Western  firebrick. 

Fireclays  are  used  not  only  for  firebrick  and  tile,  but  also  for  muffles, 
retorts,  and  clay  vessels  of  different  sorts.  The  clay  varies  much  in  plas- 
ticity. Clay  alone  is  unsuited  for  brick,  since  in  burning  it  shrinks  and 
cracks.  Firebrick  manufacturers,  therefore,  employ  a  mixture  of  one  or 
more  grades  of  clay,  adding  also  a  certain  percentage  of  coarsely  ground 
firebrick  called  "  chamotte."  The  addition  of  this  unshrinking  material 
prevents  the  cracking  that  otherwise  would  result.  The  assayer,  who 
uses  clay  for  luting,  mixes  with  it  for  the  same  reason  at  least  half  its  weight 
of  sand.  In  the  manufacture  of  firebrick  the  required  mixture  is  ground 
in  a  dry-pan,  a  machine  similar  in  construction  to  the  Carlin  mixing  pan 
(see  Fig.  170),  but  provided  with  a  bottom  made  of  perforated  plates  to 
discharge  the  material  when  ground  sufficiently  fine.  Scrapers  carried  in 
front  of  the  rollers  throw  material  in  their  path,  and  the  mixture  when 
ground  is  screened,  and  further  mixed  in  a  horizontal  pug-mill,  being  there 
tempered  by  the  addition  of  water  to  the  desired  consistence. 

The  molding  of  brick  is  done  by  hand  or  by  machine.     If  by  hand  the 
mixture  is  brought  to  the  consistence  of  mud,  and  made  into  balls  suf- 
ficiently large  to  fill  a  mold.     (See 
Fig.  20.)     The  mold  is  first  sanded 
to  prevent  the  adhesive  mud  from 
sticking,  and  this  is  thrown  into  the 
mold  with  force,  to  fill  it  completely, 

the  excess  is  cut  off  with  a  stick  or  pIQ  20. Brick-mould. 

wire,  and  the  brick  dumped  on  a 

pallet  or  board.  The  pallets  are  placed  upon  racks,  and  air-dried  until  so 
stiff  as  to  indent  but  slightly  under  pressure  of  the  finger.  They  are  then 
put  through  a  re-pressing  machine  (Fig.  21),  where  they  are  given  their 
exact  form.  When  re-pressed,  they  are  again  placed  on  pallets  and  run 
into  a  dryer  which  is  divided  into  chambers  and  heated  by  steam,  waste 


36  REFRACTORIES 

heat  or  radiated  heat,  so  that  the  last  of  the  moisture  is  removed.  The 
bricks,  now  so  coherent  that  they  can  be  handled  with  little  damage,  are 
piled  in  open  order  in  the  kiln,  already  described  (see  Fig.  19),  and  are 
burned  at  a  temperature  between  1230  and  1390°  C.,  requiring  one  to 
three  weeks  for  this. 

In  machine  molding,  called  the  "  stiff-mud  process/'  the  clay  is  tem- 
pered with  less  water,  and  is  much  stiffer  when  molded  than  in  hand- 
molding.  The  general  form  of  the  stiff-mud  machine,  known  as  the  auger 
machine,  is  that  of  a  horizontal  cylinder,  closed  at  one  end,  and  tapering 
to  a  rectangular  outlet,  the  size  of  the  cross-section  of  the 'brick  at  the  other. 


FIG.  21. — Repressing  Machine. 

Within  the  cylinder  is  a  shaft  carrying  blades  similar  to  those  in  a  pug- 
mill,  but  at  the  end  nearest  the  die,  or  outlet,  the  blades  are  replaced  by  a 
tapering  screw.  The  tempered  clay  is  fed  into  the  cylinder  at  trie  end 
farthest  from  the  die.  It  is  mixed,  and  moved  forward  by  the  blades  until 
seized  by  the  screw  which  pushes  it  through  the  die.  The  bar  of  clay 
issuing  from  the  machine  is  received  upon  a  cutting  table  and  cut  into 
bricks  by  means  of  a  wire  frame.  The  further  treatment  of  these  bricks, 
with  the  drying,  re-pressing,  and  burning,  is  like  that  of  hand-molded 
bricks. 

Another  method  of  machine  molding  is  called  the  dry-press  process. 
In  this  method  the  mixture  of  clay  and  "  grog  "  or  coarsely  ground  brick  is 
intimately  mixed  in  a  wet-pan  with  10  per  cent  of  water,  molded  in  a  dry 


BASIN  REFRACTORIES  37 

pressing  machine,  and  is  then  sent  direct  to  the  kilns  for  burning  into 
brick.  The  expense  of  drying  is  thus  saved,  but  the  brick  is  not  of  so  good 
a  quality  as  when  otherwise  made. 

To  resist  abrasion,  firebricks  must  be  hard;  to  resist  corrosion  or  slag- 
ging, dense;  and  to  resist  high  temperature  and  sudden  changes  of  tem- 
perature, porous  and  coarse  in  texture.  We  accordingly  use  the  hard 
bricks  for  door-openings,  dense  ones  for  reverberatory  furnace  walls,  and 
the  porous  and  coarse  ones  for  the  roofs.  The  refractoriness  of  a  firebrick 
depends  on  the  quantity  of  the  fluxing  bases  (especially  alkalis)  and  silica 
contained,  and  on  the  coarseness  of  the  grain.  The  grain  depends  again 
upon  the  degree  to  which  the  "  grog  "  is  ground. 

Bone  Ash. — This  is  made  by  burning  bones,  in  a  kiln  with  an  excess  of 
air,  and  grinding  the  white  residue  to  20-mesh  size.  Organic  matter  is 
thus  removed  and  an  impure  calcium  phosphate  obtained.  Though  a 
neutral  material,  this  resists  the  action  of  litharge,  and  it  is  accordingly 
used,  not  only  in  assaying,  but  in  making  the  "  tests  "  or  movable  hearths 
of  the  English  cupelling  furnace  shown  in  Fig.  278. 

BASIC  REFRACTORIES 

Dolomite. — The  alkaline-earths,  lime  and  magnesia,  are  strong  bases 
and  are  resistant  to  basic  slags,  as  shown  later,  but  are  readily  fluxed  by 
the  silica  of  silicious  slags.  Quick-lime  is  infusible,  but  is  easily  affected 
by  the  moisture  of  the  air,  and  insufficiently  coherent  to  be  used  for  making 
basic  brick.  Dolomite  is  magnesian  limestone,  and  is  a  cheap  refractory 
material.  It  is  prepared  for  use  by  burning,  much  as  is  limestone.  The 
proportion  of  lime  to  magnesia  varies  in  dolomite,  but  the  more  magnesia 
the  better  for  use  as  a  refractory.  The  composition  of  a  typical  sample 
is  as  follows: 

Per  Cent. 

CaO 31 . 62 

MgO 20 . 19 

SiO2 1.70 

FeO 1.22 

CO, .  45.35 


100.08 

Dead-burned  dolomite,  specially  prepared,  has  of  late  been  substi- 
tuted in  part  for  magnesite  as  being  cheaper. 

Magnesite. — This  is  the  most  valuable  of  the  basic  materials.  When 
magnesium  carbonate  is  calcined  at  a  high  temperature  and  dead-burned 
to  0.5  per  cent  carbon  dioxide,  the  residue  is  practically  infusible.  It  is 
used  in  grain  form  for  furnace  linings,  or  is  manufactured  into  magnesite 
brick  for  the  same  purpose.  Magnesite  is  usually  colored  dark-brown  by 


38  REFRACTORIES 

the  presence  of  about  4  per  cent  iron  oxide.  It  is  the  presence  of  the  iron 
that  enables  it  to  bond  or  set  well  in  furnace  bottoms.  Its  main  use  is 
for  basic  open-hearth  furnaces  where  the  slag  contains  as  little  as  15  per 
cent  silica.  It  is  used  also  as  a  lining  for  forehearths  (where  it  is  in 
contact  with  low-grade  corrosive  matte),  also  in  lead,  copper,  or  other  heat- 
ing or  melting  furnaces  as  well  as  for  electric  furnaces.  The  nature  of 
the  mineral  is  shown  by  the  following  analysis: 

Per  Cent. 

CaO 1.68 

MgO 42.43 

SiO2 .' 0.92 

Fe2O3  and  A12O3 4.30 

CO2andH2O.  .  .50.41 


99.74 

Carbon  Brick. — Gas  carbon,  such  as  is  used  for  arc  lights,  is  made  into 
brick  with  a  limited  amount  of  gas-tar  and  burned  in  a  kiln.  This  brick 
has  been  found  to  be  particularly  resistant  and  refractory  in  a  reducing 
atmosphere,  as  at  the  bosh  of  an  iron-furnace. 

Other  Refractory  Materials. — A  mixture  of  portland  cement  2  parts, 
clay  1  part,  and  "  chamotte  "  or  coarsely  ground  firebrick  7  parts,  moist- 
ened and  molded  into  bricks  or  blocks,  or  used  for  patching  furnaces, 
sets  quickly  and  withstands  a  white  heat  without  disintegrating.  It  is 
easily  made  and  especially  useful  for  rapid  repairs.  Only  as  much  is  mixed 
as  is  to  be  used  at  once. 

While  common  red  bricks  are  not  refractory,  the  least  fusible  can  be 
used  in  that  part  of  the  roof  of  a  reverberatory  furnace  where  the  tem- 
perature is  not  high  or  only  at  a  red  heat.  Such  bricks  are  used  for  back- 
ing firebrick  structures.  As  a  general  rule  each  kind  of  brick  should  be 
laid  in  a  material  similar  to  that  of  which  it  is  composed.  We  should 
expect  slagging  to  take  place,  for  example,  at  joints  made  of  loam-mortar 
in  firebrick.  Such  loam,  while  cheap,  is  inferior  to  fireclay.  An  analysis 
of  good  loam  gives: 

Per  Cent. 

SiO2 80.99      | 

AUG. 9.65 

Total  impurities 4.91 

Ignition  loss 4 . 43 


99.98 


Here  we  note  that  the  fluxing  bases  rise  to  nearly  5  per  cent  while 
alumina  approaches  10  per  cent,  the  ratio  of  the  most  fusible  compound 
of  alumina  and  silica.  Where  the  fluxing  bases  rise  above  5  per  cent,  there 
is  risk  of  complete  melting  at  high  temperatures. 


CHAPTER  IV 
THE  PREPARATION  OF  ORES 

We  discuss  this  under  the  general  heads  of  Sampling,  Crushing,  and 
Grinding,  Screening  and  Classifying  and  Roasting. 

We  then  take  up  the  nature  and  operation  of  metallurgical  furnaces 
and  the  principles  of  thermo-chemistry  as  preliminary  to  the  whole  ques- 
tion of  roasting. 

PRINCIPLES  OF  SAMPLING 

Sampling  consists  in  obtaining  from  a  large  quantity  of  ore  a  small 
portion  of  a  few  ounces  for  assay.  This  must  correctly  represent  the 
entire  quantity  of  the  ore,  whether  it  be  a  few  hundred  pounds  or  thou- 
sands of  tons,  a  wagon-load  or  a  ship-load.  Often  we  have  a  lot  of  ore, 
in  which  rich  pieces  mingle  with  poorer  ones,  or  even  with  waste.  In 
sampling  we  must  take  this  variation  into  account  and  represent  each  part, 
not  only  according  to  its  value,  but  also  to  its  quantity.  Often  ore  is 
bought  or  sold  upon  the  results  of  sampling.  Thousands  of  dollars  are 
involved  and  cash  is  paid  for  ore  before  the  purchaser  has  treated  it.  In 
other  cases,  ores  taken  by  the  reduction  works  are  treated  separately,  the 
owner  receiving  whatever  is  obtained,  a  charge  being  made  to  cover  the 
cost  of  treatment  and  the  profit  to  the  reduction  works.  In  this  latter 
case  sampling  could  be  omitted.  Similarly  at  a  mill  and  mine,  operated 
in  one  interest,  the  sampling  may  be  omitted  when  considered  an  unnec- 
essary expense.  Efficiency  of  the  work  is  then  determined  by  the  assay 
of  the  tailing. 

If  a  reduction  works  is  producing  lead,  copper,  or  zinc,  in  a  form  ready 
for  market,  the  metals  do  not  necessarily  require  to  be  sampled.  When- 
ever the  precious  metals  are  also  present  in  such  quantity  as  to  pay  to 
separate  them,  however,  the  metal  is  sampled  to  learn  the  values  con- 
tained before  selling  to  the  refining  works  that  is  to  effect  the  separation. 
In  blast-furnace  treatment,  ore  and  all  other  constituents  of  the  charge 
are  sampled,  assayed,  and  analyzed.  From  the  data  thus  obtained,  the 
charge  can  be  correctly  calculated  and  proportioned. 

Not  only  is  it  necessary  to  ascertain  the  value  of  ores  and  of  metals 
that  result  from  metallurgical  operations,  but  as  well  the  value  of  the 
portions  rejected.  The  efficiency  of  the  work  of  the  metallurgist  depends 

39 


40  THE  PREPARATION  OF  ORES 

upon  thorough  extraction  from  the  parts  thrown  away.  To  be  assured 
of  this,  samples  of  slag  or  tailing  are  taken  at  frequent  intervals.  In  finding 
the  value  of  a  lot  of  ore,  we  first  weigh  the  ore,  and  base  the  assay  value 
upon  the  dry  weight.  To  do  this  we  must  determine  the  percentage  of 
moisture  contained,  as  shown  by  a  "  moisture  sample."  We  then  sample 
the  ore  regularly,  and  finally  assay  the  regular  sample.  Thus,  suppose 
we  have  a  lot  of  ore  weighing  10,800  lb.,  containing  7  per  cent  moisture 
and  by  assay  54  per  cent  lead  worth  3  cents  per  pound.  Since  the  assay 
is  made  on  the  dry  weight,  we  have,  after  deducting  moisture,  10,044  lb. 
ore  containing  5424  lb.  lead  worth,  at  3  cents  per  pound,  $162.72. 

RECEIVING,  SAMPLING,  CRUSHING,  BEDDING,  AND  STORING  ORES 

The  large  smelting  works  in  the  Rocky  Mountain  region  of  the  Western 
United  States  and  Mexico  buy  their  ores  outright  from  mine-owners  for 
treatment.  Such  works  are  called  custom  works.  A  plant  treating 
principally  ore  from  its  own  mines  is  called  a  mine  works.  In  custom 
works  all  the  kinds  of  ores,  already  enumerated,  are  sampled  and  bought 
upon  a  schedule  of  charges,  generally  established  in  advance  between 
the  works  and  the  mine-owner. 

Receiving  and  Weighing. — At  reduction  works  that  purchase  ores 
(custom  works),  the  ore  arrives  either  loose  or  in  sacks.  Whether  received 
by  wagon  or  by  car,  the  vehicle  and  ore  are  weighed  together  on  platform 
scales,  thus  finding  the  "  gross  weight."  When  the  vehicle  is  emptied, 
the  weight,  called  the  "  tare,"  is  similarly  taken.  The  difference  is  the 
/  "  net  weight,"  or  the  "  wet  weight,"  and  this  is  recorded.  When  ore 
arrives  in  sacks,  the  weight  of  the  sacks  also  is  deducted.  Often  sacked 
ore  may  be  removed  to  scales  to  be  weighed,  and  only  the  weight  of  the 
sacks  deducted,  the  difference  being  the  net  or  wet  weight.  Sacks,  if  of 
sufficient  value,  are  dried  and  returned  to  the  owner.  Railroads  often 
return  them  without  extra  charge.  Sometimes  the  sacked  ore,  if  pulver- 
ulent, rich,  or  frozen,  may  be  charged,  sack  and  all,  into  the  blast- 
furnace, the  sack  serving  to  retain  the  fine  contents  until  smelted,  thus 
preventing  the  loss  of  flue  dust. 

The  Moisture  Sample. — In  theory  the  moisture  sample  should  be 
taken  at  the  instant  of  weighing,  since  the  ore  may  dry  and  become  lighter. 
The  sample  is  taken  while  the  car  is  being  unloaded  or  immediately  after- 
ward. To  represent  by  the  sample  the  ore  as  contained  in  the  car,  holes 
are  dug  at  average  points  (setting  aside  the  dry  top  layer)  and  small  por- 
tions are  taken  of  ore  that  appears  to  be  of  average  moisture.  These  por- 
tions are  put  in  a  covered  can,  and  50  oz.  of  the  mixture  are  weighed  on  a 
moisture-scale.  After  cautiously  drying  on  a  hot-plate,  or  preferably 
over-night  on  steam-coils,  the  50-oz.  portion  is  again  weighed,  and  the 


HAND  SAMPLING  41 

percentage  of  moisture  determined  by  the  loss  in  weight.     The  shipper 
often  sends  a  representative  to  watch  the  sampling  of  his  ore.     Such  a  i 
man  should  pay  attention  to  this  detail,  otherwise  too  high  a  percentage  / 
may  be  deducted  for  moisture. 

Sampling  methods  may  be  divided  into  two  classes:  hand-sampling 
and  machine-  or  automatic-sampling.  Any  method  of  sampling  includes 
the  starting  and  finishing  operations. 

Hand  Sampling. — This  includes  the  methods  called  "  grab  sampling  " 
and  "  trench  sampling,"  which  are  imperfect,  and  the  regular  methods 
known  as,  "  coning  and  quartering,"  "  fractional  selection  "  and  sampling 
with  the  "  split  shovel." 

For  determining  the  contents  of  fluxes  and  fuels  and  certain  furnace 
products,  the  "  grab  sample  "  may  serve.  It  consists  in  taking  at  uniform 
distances  over  the  pile  or  lot,  similar  amounts  broken  from  the  lumps  and 
taken  from  the  fine.  These  portions  are  mixed  and  sampled  by  coning 
and  quartering,  by  fractional  selection,  or  by  using  the  split  shovel. 

Coning  and  Quartering. — The  ore  is  crushed  and  put  in  a  circle  or  ring 
about  8  ft.  diameter  on  the  sampling  floor.  The  workman  circles  within 
this  ring,  shoveling  all  the  ore  to  the  apex  of  a  cone  at  the  center.  This 
completed  with  the  shovel  working  from  the  apex  radially,  the  ore  is 
drawn  into  a  flat  disk.  This  is  marked  by  diametral  lines  into  four  equal 
portions,  of  which  two  opposite  ones 
are  left  as  I  and  III,  Fig.  22,  on  the 
floor  and  II  and  IV  removed.  The 
reserved  sectors  are  again  shoveled 

into  a  ring,  then  made  into  a  cone, 

,     ,P    ..        .          .    , ,       r.  FIG.  22. — Quartering  an  Ore  in  Sampling, 

now  half  the  size  of  the  first  one. 

This  is  again  flattened  into  a  disk,  quartered  and  the  two  opposite  quarters 
reserved.  The  process  goes  on  in  this  way  until  the  sample  has  become  of 
small  bulk,  say  of  2  Ib.  weight,  when  it  should  be  again  ground  to  pass 
through  an  80-mesh  screen.  It  is  thoroughly  mixed  by  "  rolling  "  on  a 
sheet  of  thin  rubber  cloth,  and  the  mixed  product  distributed  into  one  or 
several  4-oz.  bottles  or  into  manila  sample-sacks  which  are  marked  with 
the  name  and  particulars  of  the  sampled  lot. 

Fractional  selection  differs  from  the  quartering  method  in  that  every 
second  or  fourth  shovelful  is  reserved  and  coned  as  above  described,  for 
the  purpose  of  mixing.  From  this  cone  each  second  or  fourth  shovelful  is 
again  reserved  and  coned,  and  this  continues  until  it  is  necessary  to  recrush. 
After  this,  reduction  in  bulk  again  proceeds  using  a  smaller  shovel,  accord- 
ing in  size  with  that  of  the  sample. 

In  sampling  by  the  split-shovel,  a  good  tool  is  shown  in  Fig.  23,  the  Brun- 
ton  quartering  shovel,  the  central  compartment  holding  the  sample.  From 
the  already-mixed  pile  shovelfuls  are  taken,  and  by  backward  movements 


42 


THE  PREPARATION  OF  ORES 


FIG.  23. — Brunton's  Quartering  Shovel. 

three-fourths  of  it  slides  from  the  shovel  blade  into  a  heap,  the  remaining 
fourth  in  the  central  compartment  being  thrown  into  a  separate  pile, 

This  pile  is  attacked  in  the  same 
way,  so  that  successively  the 
amount  to  be  sampled  rapidly 
decreases. 

The  ore  collected  upon  the 
sampling  room  floor  is  there  cu1 
down  by  cutting  and  coning  01 
by  means  of  a  Jones  sampler,  Fig 
24,  until  it  weighs  10  Ib. 

Fig.  24  is  a  view  of  a  rifflec 
sampler.     The  ore,  evenly  spreac 
in  the  scoop,  is  so  distributed  ir 
FIG.  24.— Jones  Sampler.  the  riffles  that  one-half   goes   tc 

the  right-hand  part  and  half  tc 
the  left-hand  one.  Either  half  is 
again  cut  down  in  the  same  manner, 
and  so  until  but  a  small  bulk  re- 
mains, truly  representative  of  the 
original. 

Machine  or  Automatic  Sampling. 
— It  will  be  seen  that  the  methods 
of  sampling  by  hand  as  just  de- 
scribed, especially  for  large  lots, 
involve  much  labor,  and  it  has  been 
sought  to  overcome  this  by  the  use 
of  machinery.  A  sample  from  a 
stream  of  ore,  coming  from  a  crush- 
ing machine,  and  called  a  "  running 
sample,"  is  taken  automatically, 
this  stream  being  deflected  to  one 
chute  four-fifths  of  the  time  and  to 


FIG.  25. — A  Sample-grinding  Mill. 


another,  as  a  sample,  one-fifth  of  the  time,  as  shown  in  the  Vezin  sam- 


SAMPLING  MILL 


43 


FIG.  26. — Braun  Disk  Grinder. 


pier,  Fig.  27.  It  consists  of  a  tube  carried  by  a  vertical  shaft  making  30 
R.P.M.  Attached  to  the  side  of  the  tube  and  opening  into  it  is  a  scoop. 
As  the  shaft  revolves 
"  counter-clockwise  "  the 
scoop  (occupying  one-fifth 
of  the  circumference)  cuts 
through  the  stream  of  ore 
from  the  inclined  feed- 
chute  for  one-fifth  of  the 
time.  The  ore  thus  inter- 
cepted falls  through  the 
tube  and  becomes  the 
sample,  while  the  four- 
fifths,  the  rejected  portion, 
falling  into  the  main  hop- 
per, is  delivered-  by  the 
chute  to  a  bin. 

Sampling  Mill. — Fig.  28  is  a  sectional  elevation  of  a  sampling  mill. 
An  ore  dump-car  at  the  right  discharges  its  load  into  a  sloping-bottom  ore 
bin,  whence  it  is  drawn  off  by  a  sliding  bin-gate  and  fed  to  a  15  by  9  in. 
Blake  crusher  (see  Fig.  34).  The  discharge  from  the  crusher  falls  into  the 

boot  of  a  vertical  elevator,  which  raises 
it  to  the  top  of  the  building  and  feeds  it 
to  the  chute  of  a  Snyder  sampler  or  a 
Vezin  Sampler,  Fig.  27,  where  20  per 
cent  of  the  stream  is  cut  out  to  go  to  the 
7  by  9  in.  Blake  crusher  just  below.  The 
crushed  product  of  the  crusher  is  again 
spouted  to  a  second  sampler,  where  again 
20  per  cent  of  the  stream  is  saved,  and 
thence  to  rolls  (see  Fig.  43).  The  roll- 
product  passes  on  to  a  third  sampler, 
where  again  one-fifth  or  20  per  cent  of 
the  flow  is  caught  to  be  more  finely 
crushed,  to,  say,  J  in.  in  size,  in  a  smaller 
roll,  whose  whole  product  falls  upon  the 
floor  of  the  sampling  room.  The  rejected  ore  from  upper  sampler, 
amounting  to  80  per  cent  of  the  whole,  is  shot  back  into  an  ore  dump- 
car.  The  rejected  portions  from  No.  2  and  No.  3  sampler  pass  out  at 
the  side  of  the  b.uilding  to  fall  in  a  pile  upon  the  ground. 

The  product  is  mixed  on  a  mixing  cloth,  cut  down  by  a  smaller  Jones 
riffle  to  2  lb.,  dried  in  a  steam-oven,  and  finely  ground  on  a  bucking-plate 
or  in  a  disk-grinder,  Fig.  26,  to  80-mesh  or  finer. 


FIG.  27. — Vezin  Sampler. 


44 


THE  PREPARATION  OF  ORES 


The  Braun  Disk-grinder  is  here  shown  as  opened  for  cleaning.     Fine 
grinding  is  done  between  the  two  fluted  disks  when  the  machine  is  closec 


FIG.  28.— A  Sampling  Mill. 


for  action.     The  hinged  disk  is  pressed  against  the  revolving  one,  the  flutes 
conducting  the  ore  downward  between  the  grinding  surfaces  to  the  drawei 


MACHINE  SAMPLING  45 

below.  The  ground  material  is  passed  through  a  screen,  the  oversize 
being  returned  to  the  grinder. 

Finishing  the  Sample. — The  ground  product  is  now  mixed  by  "  rolling  " 
on  a  rubber  mixing  cloth  and  distributed  into  manila  paper  sample-sacks 
that  hold  3  or  4  oz.  each. 

The  final  operations,  from  the  taking  from  the  sample  safe,  are  called 
"  finishing  the  sample." 

At  a  custom  works  all  the  ore  is  stored  in  a  sample  safe  or  can  and  there 
is  held  until  the  ore-lot  has  become  the  property  of  the  works  by  purchase. 

Sampling  of  Ores  Containing  Metallic  Substances. — This  is  an  opera- 
tion requiring  a  clear  knowledge  of  the  principles  of  sampling.  We  come 
upon  these  "  me  tallies  "  sometimes  in  the  operation  of  sampling.  They 
must  be  separated,  cut  smaller,  and  quartered  down  separately  by  a  hand- 
method,  and  reduced  in  size,  at  the  same  rate  as  the  fine  ore.  If  a  fine  sub- 
stance is  made  by  cutting  up  the  metallics  it  can  be  united  with  fine  ore. 
Often  metallics  are  brittle,  but  with  diligent  work  can  be  broken,  cut  and 
"  quartered  down  "  without  serious  difficulty. 

Cost  of  Sampling. — In  1910  the  cost  of  moving  the  ore  cars,  unloading 
into  bins,  returning  the  cars  to  the  sampling-mill,  and  unloading  the  frac- 
tional part,  usually  one-tenth,  retained  was  taken  at  10  cents  per  ton.  The 
cost  of  hand-sampling  the  tenth  part  was  taken  at  75  cents  per  ton. 
Hence,  for  unloading  and  hand-sampling  a  100-ton  lot,  the  total  cost  was 
17.5  cents  per  ton.  At  the  Metallic  Extraction  Works,  Cyanide,  Colorado, 
ore  was  then  unloaded  from  the  car  to  a  feed-chute  crushed  to  J-in.  size, 
automatically  sampled  and  delivered  to  storage  bins,  for  1 1  cents  per  ton. 
A  charge  of  $1  to  $2  per  ton  has  been  made  for  sampling,  storing,  assaying, 
and  selling  ore  at  custom  works  or  sampling-mills,  where  the  company  has 
acted  as  selling  agent  and  obtained  the  best  possible  price  for  the  shipper. 
The  price  for  sampling  concentrate  was  50  cents  per  ton  less. 

Sampling  Concentrate,  Tailing,  and  Ore-pulps. — Concentrate  is 
sampled  easily,  for  it  can  be  thoroughly  mixed  and  sampled  by  hand. 
Tailing  carries  40  to  50  per  cent  moisture.  It  has  but  little  value,  and 
needs  no  close  attention.  Ore-pulp  flowing  in  a  launder  is  often  auto- 
matically sampled. 

Fig.  29  shows  the  machine  used  for  this.  The  pulp  flow,  entering  by 
the  chute  on  the  left,  is  carried  away  by  a  launder  set  to  receive  it.  As 
the  wheel  revolves,  the  attached  scoops  divert  part  of  the  flow  into  a 
compartment  of  the  receiving  box.  When  not  so  sampled,  a  bucketful, 
taken  each  hour  from  the  stream,  and  all  these  samples  united  in  one  por- 
tion after  decanting  the  water,  may  be  used  to  determine  the  approximate 
daily  average  value.  When  loaded  into  cars,  the  sample  is  sometimes 
taken  by  boring  to  the  bottom  of  the  body  of  ore,  using  a  ship  auger,  or 
again  by  a  pipe  driven  downward.  This  is  a  kind  of  grab-sampling,  but 


46 


THE  PREPARATION  OF  ORES 


serves  in  mills  where  such  an  approximation  is  considered  sufficiently 

accurate. 

Sampling  Iron  Ores. — These,  being  uniform  in  constitution,  are  more 

simply  sampled.  The  railroad  cars 
are  sampled  by  taking  a  grab- 
sample  at  six  or  eight  places  uni- 
formly over  the  load.  From  this 
an  analysis  determines  the  char- 
acter of  the  ore,  and  where  it  is  to 
be  stored  for  shipment  by  ore-boat. 
Later  the  cargo  is  sampled  at  the 
receiving  port.  This  is  done  by 
taking  grab-samples  upon  the  ex- 
posed surfaces  of  the  ore  in  the 
hatches  while  it  is  being  unloaded 
the  ore  being  immediately  put  in 
closely  covered  cans  so  that  its 
moisture  shall  be  conserved. 

Mill  Samples. — The  value  of 
the  ore  going  to  the  mill  (the  mill 
heads)  is  estimated  as  the  value  of 
the  product  plus  the  tailings  loss. 
As  a  check  a  sample  is  taken  of 

FIG.  29.-Tailings  Sampler.  the  ball-mi11  discharge,  this  being 

dried,  solution  and  all  and  assayed. 

Knowing  the  specific  gravity  of  the  discharge  we  can  compute  its  con- 
tained solution,  whose  value  is  subtracted  from  that  of  the  solution 
taken  as  a  drip  from  the  storage  solution  tank.  This  checks  within  4  per 
cent  of  the  product-plus-tailings  result. 

Dip-samples  are  taken  hourly  of  the  various  agitator  and  thickener 
pulps  and  solutions.  To  prevent  further  dissolution  by  the  cyanide,  10 
c.c.  of  a  10  per  cent  solution  of  sodium  sulphide  is  added  to  each  dip- 
solution  once  per  shift.  They  are  filtered,  washed  and  dried  for  assay, 
and  results  promptly  reported. 

Principles  of  Sampling. — In  the  progressive  crushing  above  described, 
it  will  be  observed  that  the  ore  is  made  finer  as  the  sample  becomes  less. 
This  is  to  make  sure  of  a  constant  ratio  between  the  size  of  a  single  rich 
piece  and  the  whole  sample,  that  such  rich  piece  shall  not  produce  an 
appreciable  effect  on  the  assay  value,  whether  it  be  present  or  absent. 
The  richer,  and  at  the  same  time  the  more  "  spotty  "  or  varied  the  ore, 
the  finer  it  should  be  crushed  before  cutting-down  or  quartering.  The 
table  below  shows  how  this  is  arranged  in  practice: 


MACHINE  SAMPLING 


47 


SIZE  OF  LARGEST  PIECES 


VALUE  IN 

SILVER  OUNCES 

PER  TON. 

Weight  of  Ore. 

Highest  300, 
Average  50. 

Highest  3000, 
Average  75 

Highest  10,000, 
Ave.  500. 

100  tons  to  10  tons 

Cocoanut 

Fist 

Fist 

10  tons  to  1  ton  

Orange 

Egg 

Walnut 

1  ton  to  200  Ib  

Walnut 

Chestnut 

Chestnut 

200  Ib  to  5  Ib 

Pea 

Wheat 

Wheat 

5  Ib  to  bottle  sample 

20-mesh 

25-mesh 

50-mesh 

Bottle  sample  

80-mesh 

100-mesh 

120-mesh 

We  may  conclude  that  for  accurate  sampling  the  requirements  are: 

(1)  The  taking  uniformly  frequent  portions  to  ensure  an  average  of 
the  stream  of  ore  as  it  is  undergoing  progressive  sampling. 

(2)  Thorough  mixing  of  the  ore  to  ensure  uniform  richness. 

The  Martin  Sampling  Machine. — This  machine  excavates  the  slime  or 
mud  of  flotation  concentrates  containing  20  per  cent  moisture  from  rail- 
road cars  transferring  it  to  bins  and  taking  out  a  sample  in  so  doing. 

The  bin,  long  and  narrow,  has  a  slit  4  ft.  wide  in  the  bottom  covered  by 
transverse  plank.  The  furnace  charge-car  runs  beneath  the  bin,  so  that 
the  material  is  drawn  off  to  the  car  by  taking  up  the  planks  progressively. 
Centrally  above  the  bins  is  the  sampling  machine  which  travels  freely 
delivering  the  slime  where  desired. 

On  the  nearby  parallel  track  is  stationed  the  train  of  cars  loaded  with 
the  slime.  The  machine  is  furnished  with  a  grab-bucket,  and  moves  along 
the  track  where  needed  for  excavating.  The  load  is  delivered  upon  a 
slowly  moving  8-ft.-wide  endless  sheet-metal  belt  discharging  thence  upon 
a  reel  the  width  of  the  belt.  The  reel  is  made  up  of  several  disks  18  in. 
diameter  with  wires  stretched  from  end  to  end  over  them  at  6  in.  intervals. 
The  sheet  of  slime  is  chopped  up  by  the  disks  and  wires  of  the  reel  and  falls 
in  lumps  upon  a  hollow  cylinder  of  15-in.  diameter  having  a  slit  in  it  of 
one-tenth  its  circumference.  As  the  cylinder  revolves  the  material  falls 
upon  it,  and  when  the  slit  comes  under  them  it  falls  out  of  the  slit  again 
upon  a  conveying  belt  covered  with  a  dry  calcareous  sand  from  a  nearby 
deposit.  This  belt  takes  this  supply  from  a  bin,  the  head  pulley  of  it  being 
within  the  traveling  sample  mill  and  delivering  its  load  of  sand  and  slime 
where  another  divider  belt  takes  out  a  tenth,  so  that  a  1  per  cent  sample  is 
obtained. 

It  will  unload  80  tons  per  hour,  where,  when  unloading  by  hand,  the 
cost  was  20  cents  per  ton. 

Pulp  Sampler. — As  shown  in  Fig.  30,  this  consists  of  a  disk  with  an 
attached  pipe.  The  disk  is  counter-weighted,  so  that  when  past  the  axis, 


48 


THE  PREPARATION  OF  ORES 


it  suddenly  falls,  the  pipe  sweeping  through  the  tailings  stream.     One  may 
note  a  slot  in  the  pipe  where  the  pulp  from  the  tailings  stream  enters.     As 


O-  Section  of  Pipe 


,  Pipe  Cap 


form  Gear 


Wieght 


Collar  tight 
ti  Shaft 


FIG.  30. — Pulp-sampler. 

the  pipe  is  lifted  with  the  disk  again  into  its  vertical  position  the  pulp 
flows  out  of  the  lower  end  of  the  pipe  into  the  sample  launder. 

SAMPLING  METALS 

Metals  may  be  sampled  either  in  the  solid  or  molten  state. 

Gold  or  Silver  Bars  or  Ingots. — These  are  sampled  for  assay  either  by 
granulating  a  small  portion  of  them  or  by  taking  chip-samples  from  them. 
In  the  first  case,  while  the  metal  is  in  a  molten  condition,  a  small  ladleful, 
weighing  an  ounce  or  less,  is  taken  from  the  crucible  immediately  after 
stirring  it.  This  is  poured  into  a  bucketful  of  water,  thereby  granulating 
the  metal  and  forming  particles  of  a  variety  of  sizes,  convenient  for  weighing 
and  assay.  Chip-samples  are  taken  at  points  diagonally  opposite -on  the 
edges  of  the  bar,  and  a  cold-chisel,  cutting  out  a  small  wedge-shaped  piece, 
is  used  for  this  purpose.  The  pieces  are  annealed  and  rolled  into  a  ribbon 
for  assay.  The  average  assay-value  of  the  two  pieces  thus  obtained  is 
taken  as  the  true  value. 

Base-bullion. — This  is  lead  that  comes  from  silver-lead  blast-furnaces, 
and  it  contains  commonly  100  to  400  oz.  of  silver  per  ton.  When  poured 
into  molds  to  be  cast  in  bars,  the  silver  segregates,  and  the  exterior  of  the 
bar,  that  cools  first,  is  richer  in  silver  by  several  ounces  than  the  central 
part.  This  is  illustrated  in  the  cross-section  of  a  bar  (Fig.  31),  in  which 


SAMPLING  METALS 


49 


the  center  of  the  bar  assays  10  ounces  less  than  the  exterior, 
is  sometimes  sampled  by  taking  two  "  chips  "  or 
punchings,  one  from  the  top  and  one  from  the  bottom 
of  each  bar.  The  punch  (Fig.  32),  resembling  a  belt 
punch,  is  8  in.  long  and  removes  a  cylindrical  piece  of 
1|  in.  diameter  by  about  \\  in.  length. 

From  a  carload  lot  of  400  bars,  800  of  these  chips 
would  be  obtained.     These  are  melted  and  the  fused 


Base-bullion 


«  260.fi  2G4.1 


.6  250.0  219.0  219.0 


•r>  257.0  242.0  258.0 


2ol.O  259.5 


239.0  2i,0. 


261.5 


Average  258.2  oz.  silver  per  1 


of  Silver  in  a  Bar  ot 
Base-bullion. 


metal  stirred  in  a  plumbago  crucible   and   cast  into   FIG>  31. Distribution 

bar.     This  is  a  sample  of  a  400-bar  lot  of  base-bullion. 

equivalent  to  20  tons.     A  better  way  of  sampling, 

however,  is  to  remelt  the  metal  in  a  large  kettle  (see 

Fig.  276),  and  to  skim  and  recast  into  bars  for  shipment.  While  casting 

the  metal,  a  sample  is  taken  from 
the  molten  bath  and  poured  into  a 

bullet-mold  of  such  a  size  that  each 
FIG.  32. — Base-bullion  Sampling-punch.     •_'  «  .         •  ,  •       ,   i  u   ir 

bullet  weighs  approximately  a  half 

assay-ton.     This  is  trimmed  to  the  exact  weight  for  the  assay. 

Copper  Ingots  or  Anodes. — The  segregation  of  gold  and  silver  in  copper 
ingots  is  even  more  marked  than  in  bars  of  base-bullion.  This  is  shown 
in  Fig.  33,  which  represents  the  distribution  of  gold  and  silver  in  an  ingot 
of  blister-copper  5  in.  deep.  In  this 
case,  however,  the  interior  is  higher 
both  in  silver  and  gold.  The  usual 
way  to  sample  such  bars  is  to  drill  into 
them  and  retain  the  borings  for  a  sam- 
ple. Manifestly  a  sample  like  this  is 
uncertain,  and  depends  upon  the  selec- 
tion of  the  place  on  the  ingot  for  taking 
it.  To  obviate  this  difficulty,  in  sam- 
pling a  lot,  say  of  100  bars,  it  is  custom- 
ary to  drill  into  each  succeeding  bar  at  a  different  spot  to  obtain 
an  average  by  so  doing.  It  is  preferable,  however,  to  take  the  sample  at 
the  time  the  copper  is  melted  and  well  mixed  in  the  furnace  by  poling. 
As  in  the  case  of  base-bullion,  samples  of  copper  are  taken  while  dipping 
or  casting,  one  at  the  beginning,  one  when  the  charge  is  half  removed,  and 
one  toward  the  end.  The  average  of  the  three  samples  is  regarded  a  cor- 
rect representation. 

Pig  Iron. — This  is  sampled  and  graded  by  inspecting  its  fracture  and  by 
chemical  analysis.  When  an  analysis  is  to  be  made,  the  sample  is  taken 
from  the  drillings  of  a  small  bar,  molded  while  the  metal  is  flowing  from 
the  furnace.  The  percentage  of  silicon  determines  the  grade,  the  whole 
of  the  sample  pieces  being  dissolved  for  assay. 


1 192.1  j  195.2  i  194.5 1 122.0 1  68.1 
I  0.321  0.34'  0.34J  0.26"  0.20, 

|il4.5|ll7.0lll2.3|105.r|  5.72 
I  0.221  «.28j_0.30|_  0.26^0.20 
P71.3i  70.31  6U.8'  6\>£\  70.5 
j  0.24]  0.22  j  0.22]  0.22]  0.22 


FIG.  33.— Section  of  Bar  of  Ingot 
Copper. 


CHAPTER  V 
CRUSHING  AND    GRINDING,    SCREENING    AND    CLASSIFYING 

Size  of  Run-of-mine  Ore. — This,  as  it  comes  from  underground,  or 
I  from  open-cut,  will  vary  from  pieces  several  tons  in  weight  to  dust  which, 
I  when  wet,  is  called  slime.     The  larger  pieces  may  be  broken  to  sizes  suited 
to  loading  by  hand  sledging  or  by  block-holing,  that  is,  by  drilling  and 
shooting  them.     Since  this  is  more  expensive  than  breaking  by  rock- 
breakers,  the  plan  is  to  have  these  machines  large  enough  to  take  such 
pieces,  especially  if  they  are  hard  to  break. 

PRINCIPLES  OF  CRUSHING 

Power  Needed  in  Crushing. — The  work  done  in  crushing  varies  inversely 
with  the  size  to  which  the  ore  particles  are  crushed,  or  directly  according 
to  the  increase  of  surface. 

Thus  a  ton  of  quartz  ore,  composed  of  1-in.  cubes,  if  crushed  to  0.5-in. 

cube  would  need  0.257  horse-power  to  so  reduce  it.     The  diameter  of  these 

I    smaller  cubes  will  then  be  one-half,  while  their  surface  will  be  double  that 

of  the  original  ones.     The  surface  of  a  1-in.  cube  is  6  sq.  in.,  and  that  of 

.   the  resultant  eight  0.5-in.  cubes  will  be  8X6X0.52=12  sq.  in.     Were  the 

1-in.  cubes  crushed  to  0.25  in.,  or  one-fourth,  the  surface  of  them  would 

be  24  sq.  in.  and  the  power  needed  would  be  0.514  H.P.     This  rule  is  fairly 

constant  with  the  coarse  sizes,  but  with  the  finer  ones  the  increase  of  surface 

is  more  rapid  than  the  work  needed  to  produce  them. 

Action  of  Machines  in  Crushing. — Crushing  may  be  effected  in  either 
of  two  ways :  In  the  first,  the  breaking  is  due  to  the  impact  of  two  approach- 
ing surfaces,  as  in  rock-breakers,  stamps,  or  rolls.  In  the  second  the  sur- 
faces move  over  one  another,  the  ore  being  interposed  and  abraded  or 
sheared,  as,  for  example,  in  grinding  pans,  coffee-mill  grinders,  and  the 
Braun  sample-grinder.  Often  both  actions  are  taking  place,  as  in  ball- 
mills  and  tube-mills. 

Stage  Grinding. — The  operations  of  reducing  an  ore  to  a  fineness,  such 

that  its  mineral  particles  can  be  acted  on  by  a  solvent,  separated  by 

concentration  or  roasted,  is  done  by  crushing  in  stages.     The  first  stage, 

or  coarse-crushing,  is  done  by  rock-breakers  reducing  the  ore,  so  that  its 

1  larger  piece  will  be  no  more  than  1.0  to  2.0  in  diameter.     The  second,  or 

50 


COARSE  CRUSHING  5J 

intermediate  crushing,  aims  to  reduce  the  coarse-crushed  ore  to  30-  or  40- 
mesh  size,  this  being  done  by  stamps,  rolls,  or  ball-mills.  The  third,  or 
fine  grinding,  takes  this  product  and  grinds  it,  so  that  much  of  it  will  pass 
a  200-mesh  screen. 

A  point  to  be  observed  in  preparing  ore  for  leaching  is  to  avoid  making 
slime  in  crushing.  Any  considerable  portion  of  finely  ground  or  slimed 
material  hinders  percolation  greatly.  If  grinding  is  performed  by  rolls,  a 
more  granular  product,  containing  less  slime,  isjgreduced  than  by  crushing 
to  the  same  screen-size  with  stamps.  If  it  is  desired  to  obtain  the  maxi- 
mum quantity  of  sand  and  the  minimum  of  slime,  then  gradual  reduction, 
or^raded_crushing,  should  be  adopted.  This  consists  hi  first  screening 
out  the  ore,  already  sufficiently  fine. 

Screen  Sizes. — We  specify  the  size  of  a  piece  of  ore  by  saying  that  it 
will  just  pass  through  a  2-in.  or  4-in.  ring,  for  instance,  or  through  a  screen 
having  round  or  square  openings  of  that  size.  For  smaller  pieces  or.  particles, 
the  size  is  designated  as  being  able  to  pass  through  a  wire-mesh  screen,  the 
distance,  center  to  center  of  the  wire,  being  meant)  The  opening  is  less 
than  this  distance  by  the  thickness  of  the  wire.  Thus  a  20-in.  mesh  screen 
would  be  one  having  20  wires  to  the  inch,  or  0.05  in.  center  to  center,  and 
also  20  openings,  whose  size  would  depend  on  the  thickness  of  the  wire. 
Thus,  the  diameter  of  the  wire  might  vary  from  0.025  to  0.009  in.,  and  the 
resultant  opening  from  0.025  to  0.041  in.  Where  the  thickness  of  the  wire 
is  equal  to  one-half  the  mesh,  then  the  percentage  of  opening  is  25  per  cent, 
and  the  largest  particle  that  can  pass  the  screen  is  -jV,  or  0.025  in.  This 
proportion  is  sanctioned  when  the  actual  wire  size  has  not  been  specified. 

COARSE  OR  PRIMARY  CRUSHING 

Two  kinds  of  rock-breakers  are  in  general  use  for  coarse  crushing — 
the  Blake  and  the  gyratory.  Blake  crushers,  with  a  feed-aperture  of  5  ft. 
by  6  ft.,  have  been  made  so  that  pieces  of  rock  of  nearly  that  size,  and 
weighing  four  or  five  tons,  can  be  broken.  Such  large  machines  are  expen 
sive,  and  so  hand-breaking  may  be  done  when  the  rock  favors  it.  Blake 
crushers  have  been  made  to  take  a  boulder  as  large  as  5  ft.  while  recently, 
gyratory  crushers  have  been  built  having  a  receiving  opening  of  5X15  ft., 
weighing  237  tons,  and  having  a  capacity  of  2500  tons  per  hour.  For  the 
same  width  of  aperture  a  gyratory  crusher  costs  nearly  two  and  a  half  times 
the  Blake.  It  is  convenient  to  remember  that  either  kind  requires  about 
one  horse-power  to  crush  one  ton  per  hour. 

The  Blake  Crusher. — Fig.  34  is  a  view  of  a  Blake  crusher  or  breaker. 
Fig.  35  is  a  section  of  one  having  a  receiving  opening  20X20  in.  This 
crushes  ore  as  it  comes  from  the  mine,  containing  pieces  as  large  as  12  in. 
diameter,  at  the  rate  of  25  tons  per  hour.  There  are,  however,  clayey 


52 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


and  talcose  wet  ores  containing  25  to  30  per  cent  moisture  that  stick  to 
the  rock-breaker,  and  are  impossible  to  crush  in  the  wet  state.  Such  ore 
may  be  first  dried  in  a  cylindrical  dryer.  The  Blake  rock-crusher  is  shown 
in  perspective  in  Fig.  34  and  in  longitudinal  section  in  Fig.  35.  It  con- 
sists of  a  heavy  cast-iron  frame,  marked  N,  within  which  is  placed  the  fixed 
jaw  F  and  the  swinging  jaw  B  and  between  them  the  ore  is  crushed.  A 
shaft  T,  eccentric  where  it  passes  through  the  pitman  K,  causes  this  to 
rise  and  fall,  producing  a  corresponding  movement  of  the  adjacent  ends  of 
the  toggles  J,  J.  As  these  rise,  the  effect  is  to  push  the  jaw  forward  to 


FIG.  34. — The  Blake  Ore-crusher. 


produce  the  crushing  movement.  As  the  pitman  and  the  toggles  Descend 
the  jaw  recedes,  and  is  pulled  back  by  the  spring  rod  P  and  the  spring  R. 
Flywheels  A  help  to  steady  the  movement.  The  machine  is  driven  by 
the  pulley  at  250  revolutions  per  minute.  The  movement  of  the  lower 
end  of  the  jaw  is  \  to  f  in.  For  the  breaker  above  specified,  the  discharge- 
opening  would  be  20  by  1J  in.  to  crush  to  IJ-in.  size.  The  receiving  opening 
would  be  20  by  12  in.,  and  would  take  pieces  as  large  as  12-in.  diameter. 

The  Gyratory  Crusher. — Fig.  36  is  a  perspective  view,  and  Fig.  37 
a  sectional  elevation  of  this  type  of  crusher.  Referring  to  Fig.  37  (the 
bottom  plate  dropped)  is  a  main  frame  or  body  2  and  3,  with  a  three- 


COARSE  CRUSHING 


53 


FIG.  35. — Blake  Ore-crusher  (section). 


FIG.  36. — Gyratory  Crusher. 


54  CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


FIG.  37. — Gyratory  Crusher  (section). 


FIG.  38.— Overflow  Ball-mill  (section). 


COAHSE  GRINDING  55 

legged  spider  to  carry  the  top  of  the  spindle  or  vertical  shaft.  This  is  sur- 
mounted by  an  ore-hopper  7,  the  ore  falling  between  the  legs  of  the  spider 
to  be  crushed  between  the  liners,  19  of  the  body  3,  and  the  cone-shaped 
head  of  the  spindle  25.  The  lower  end  of  the  spindle  is  moved  in  a  circle 
without  revolving,  by  an  eccentric  sleeve  8,  made  in  one  with  the  bevel 
gear  9.  Thus  the  opening  between  the  head  and  the  liner  is  alternately 
opened  and  closed,  crushing  the  ore,  the  product  discharging  over  the 
chilled  wearing-plates  22.  The  amount  of  the  jaw-opening  can  be  varied 
by  raising  the  spindle  using  the  lighter-screw  29.  At  12  is  the  belt- 
driving  pulley. 

INTERMEDIATE  OR  FINE  CRUSHING  OR  COARSE  GRINDING 

This  is  done  largely  by  stamps,  by  pans,  by  ball-mills,  by  rolls,  by 
Chilian  mills  and  by  the  Symons  disk-crusher.     Ball-mills  and  pans  are 


FIG.  39. — Marcy  Ball-mill  (perspective  view). 

good  intermediate  grinders  up  to  100-mesh,  and  to  this  point  they  are 
economical.  There  is  a  tendencyaT~present  to  supplant  stamps^  hereto- 
fore so  largely  used,  by  ball-mills,  except  in  cases  where  inside  and  outside 
amalgamation  is  to  be  used. 

The  Stamp-mill. — This  is  given  under  the  head  of  amalgamation  as 
used  in  gold-mill  practice.  It  will  take  ore  of  1  to  1$  in.  in  diameter 
and  will  reduce  it  to  pass  a  battery-screen  aperture,  economically  of 
not  less  than  0.15  to  0.4  in.  The  stamp-mill  has  reached  its  greatest 
development  on  the  Rand  in  South  Africa.  It  is  held  that  while  heavier 
stamps  have  been  used,  those  of  1600  Ib.  falling  weight  are  the  practical 


56 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


limit.     The  tendency  in  American  practice  is  to  replace  the  stamps  by 
ball-mills  in  new  construction. 

The  Ball-mill. — Fig.  38  is  a  view  of  a  ball-mill,   showing  the  spiral 
scoop  that  picks  up  the   feed    and  delivers   it  into    the  mill   and   the 


Mbtor 


der 


^Combination 
Feeder 


FIG.  40. — Ball-mill  in  Closed  Circuit. 

flaring  discharge  for  the  exit  of  the  ground  pulp.  The  discharge  is 
screened  to  remove  .the  finer  sand,  while  the  coarser,  discharging  at  the 
extreme  end,  is  returned  for  regrinding.  The  mill  has  a  self -locking  lining 
needing  no  bolts  through  the  cylindrical  shell.  The  mill  is  filled  about  half 
full  of  flint-pebbles,  such  as  are  employed  for  tube-mills,  generally  imported 

from  Scandinavian  countries  or  from 
France.  In_  place  of  pebbles,  iron 
or  steel  balls  are  increasingly  in  use, 
these  often  toughened  by  the  addi- 
tion of  manganese  or  chromium. 
The  discharge  of  the  mill  may  be 
protected  by  a  grid  or  perforated 
plate,  see  Fig.  39,  as  in  the  Marcy 
mill,  to  hold  the  balls  or  pebbles 
back  while  permitting  the  escape  of 
the  ground  material. 

Proportions  and  Efficiency. — Fig. 
41  shows  the  paths  of  travel  of 
particles  in  a  ball-mill  8  ft.  diameter 
by  6  ft.  long.  It  is  assumed  that  the 
steel  balls  are  from  3  in.  to  2  in. 
much  smaller  than  this  they  are 


FIG.  41 . — Path  of  Travel  of  Ore-particles. 
diameter,   and  that  when  they  are 


COARSE  GRINDING 


57 


removed,  as  being  a  hindrance  rather  than  an  aid  in  comminution,  being 
replaced  by  3-in.  balls.  At  the  correct  speed  22  R.P.M.  there  will  be 
needed  180  H.P. 

On  a  medium-hard  porphyry  ore,  such  a  mill  should  grind  in  twenty- 
four  hours  from  2^  in.  size  to  minus  48-mesh  about  45  tons,  or  of  a  hard 
quartz  ore  half,  and  of  soft  porphyry  twice  this.  We  distinguish  between 
.the  circulating  load,  which  is  that  which  makes  the  cycle  through  the  mill 
and  classifier,  and  the  input  load,  being  the  original  ore  plus  the  circulating, 
load.  The  water  used  will  vary  from  30  per  cent  to  50  per  cent  of  the  ore 
for  the  most  efficient  grinding.  The  mill  is  lined  throughout  with  heavy 
manganese-steel  plates. 

The  Hardinge  Conical  Mill. — This  ball-mill;  as  shown  in  Fig.  42, 
has  conical  ends  with  a  cylindrical  center  of  3.6  in.  for  a  mill  8  ft.  diameter 


FIG.  4'2. — Hardinge  Conical  Mill. 

by  22  ft.  long.  It  feeds  and  delivers  as  in  the  Marcy  mill.  Due  to  its 
shape,  the  larger  balls  keep  to  the  larger  diameter,  so  that  only  the  finely 
ground  product  and  the  worn  particles  from  the  balls  discharge.  Each 
day  a  few  new  balls  are  fed  into  the  feed-scoop. 

For  best  performance  in  a  mill  of  this  size  the  balls  would  weigh  28,000 
Ib.  and  would  occupy  0.3  to  0.4  of  the  volume  of  the  mill,  or  as  much  as  is 
shown  in  the  figure.  It  should  yield  1\  tons  per  hour  on  average  ore. 
covering  from  \  in.  to  200-mesh  size. 

Fig.  40  is  an  elevation  of  a  Marcy  ball  bill,  4J  ft.  diameter,  inclosed 
circuit  with  a  Dorr  classifier.  At  the  right  end  is  a  feeder,  which  scoops 
up  the  coarsely  crushed  ore  and  the  water  from  the  box  which  receives  this 
feed  from  a  launder.  Entering  the  mill,  the  ore  is  crushed  by  the  tough  1 
manganese  steel  balls  that  are  lifted  up  and  fall  upon  it,  due  to  the  revolu- 
tion of  the  mill.  The  mill  is  driven  by  herring-bone  gearing  and  electric 


58 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


motor  at  230  R.P.M.  The  discharge  goes  to  the  lower  end  of  the  Dorr 
classifier.  Here  the  unground  particles  sink  and  are  gradually  raked  to 
the  upper  end  to  return  to  the  mill-feed  box.  At  the  lower  end  of  the 


FIG.  43. — Fifty-four  Inch  Crushing  Rolls. 

classifier  the  floating,  finely  ground  product  or  slime  overflows,  ready  for 

further  treatment. 

Crushing   Rolls.  —  We    show  in    Fig.  43  a  perspective  view  of  rolls 

having  grooved  shells  or  tires;  in 
Fig.  44  plan  and  elevations  of  rolls 
having  smooth  shells.  In  the  plan 
and  outside  elevation  in  Fig.  45 
the  fixed  roll  is  at  the  right,  in  Fig. 
43  at  the  left,  and  the  movable  or 
spring-roll  is  carried  in  sliding  boxes. 
The  long,  heavy  tension  rods 
have  capstan  nuts  at  both  ends  for 
tightening  the  springs  against  the 
sliding  boxes.  In  case  a  hard  object 
gets  into  the  rolls  as  they  revolve, 
such  as  a  hammer-head,  the  springe 
give  back,  permitting  its  passage,  so 
that  the  rolls  do  not  stall.  In  Fig.  45 


FIG.  44.— Angle  of  Nip. 


the  roll-shell  or  tire  is  pinched  between  the  fixed  roll  head  7,  and  the  mov- 
able one  8,  due  to  a  slight  taper  on  the  heads  corresponding  to  a  like  one 
on  the  shell.  Bolts  draw  those  heads  tightly  together.  For  heavy  work 


COARSE  GRINDING 


59 


such  rolls  have  been  built  up  to  72  in.  diameter  by  24  in.  face.     They  will 
nip  even  to  4-in.  pieces,  and  with  a  "  choke  feed,"  that  is,  with  a  thick 


stream  of  ore  fed  on  to  the  rolls  while  well  set  up,  will  run  smoothly  with  a 
capacity  of  2500  tons  per  day.  Due  to  its  grooved  shell,  the  smaller 
rolls,  Fig.  43,  will  take  as  large  a  piece. 


60 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


Present  practice  in  roll  crushing  involves  a  high  peripheral  speed, 

not  only  as  increasing  capacity,  but  also  to  insure  smoother  running,  owing 

to  the  fact  that  the  inertia  of  the  rolls  carries  them  safely  by  a  sudden   • 

peak  of  load.     Such  speeds  are  from  300  to  500  ft.  peripheral  speed  or, 

i  for  42-in.  rolls,  from  27  to  45  R.P.M.     It  is  the  custom  to  drive  the  rigid 

\  roll  by  a  large  pulley,  using  a  smaller  pulley  on  tj^e-sprmg-rolL     The  fiyed 

,  or  rigid-roll  shaft  has  a  deep  groove  turned  at  one  end  of  it,  fitted  with  a 

thrust-bearing.     By  means  of  two  bolts  this  thrust-bearing  may  be  moved 

axially,  and  with  it  the  shaft.     This  results  in  giving  another  surface  of 


FIG.  46.— Chilian  Mill. 

contact  between  the  rolls,  preventing  them  from  grooving.  In  sonle  makes 
of  rolls  this  end  movement,  called  "  floating  "  is  slowly  and  automatically 
performed.  Rolls  are  also  made  without  springs;  they  are  called  "  rigid  " 
rolls.  They  have  the  advantage  that  they  produce  less  fines  and  practically 
no  oversize.  They  run  with  little  jar  or  vibration.  The  size  of  feed  for 
42-in.  rolls  would  be  1|  to  2  in.  and  the  reduction  of  the  size  of  the  feed  is 
preferably  four  to  one,  that  is  2-in.  pieces  should  be  reduced  to  \  in.  size. 

The  Chilian  Mill. — Fig.  46  is  a  view  of  a  5-ft.  mill  with  the  shrouding/ 
broken  away  to  show  two  of  the  three  rollers.     These  have  steel  tires. 
They  roll  upon  a  die-ring.     Ore  is  fed  by  launder  to  a  central  hopper  set 


COARSE  GRINDING 


61 


above  the  roll  axles,  it  flows  downward  into  the  pan  below  and  outward 
so  as  to  be  ground  between  the  rollers  and  the  ring.  It  is  splashed  outward, 
and  when  ground  fine  enough,  escapes  through  the  peripheral  screens  that 
form  the  sides  of  the  pan  and  into  a  gutter  by  which  it  flows  away.  A 
5-ft.  mill,  running  at  40  R.P.M.,  will  grind  25  to  35  tons  of  ore  in  twenty- 
four  hours,  requiring  10  H.P.  to  drive  it,  and  using  400  to  1000  gallons  of 
water  per  hour. 


FIG.  47. — Symons  Disk-crusher. 


FIG.  48. — Symons  Disk-crusher  (section). 

Besides  the  jast-running  typejof  Chilian  mill  above  described  there  is 
the  slowjirjond^called  the  Jjane^_running  12  to  15  R.P.M.  and  having 
rollers  7  ft.  diameter  by  22  in.  face,  and  weighing  7  to  8  tons  each.  These 
travel  on  a  die  ring  7  ft.  diameter  and  crush  through  a  30-mesh  screen  at 
the  rate  of  15  tons  in  twenty-four  hours.  / 

The  Symons  Vertical  Disk  Crusher. — We  show  in  Fig.  47  a  per- 
spective view;  in  Fig.  48  is  a  longitudinal  section  of  a  48-in.  Symons 
disk  grinder.  Referring  to  the  section  Fig.  48,  there  are  two  saucer- 


62 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


shaped  disks  carried  at  the  end  of  a  sleeve  and  with  a  slit  opening  revolving 
at  100  R.P.M.  Within  the  sleeve  is  a  shaft  whose  tail-end  receives  a 
reciprocating  motion,  due  to  an  eccentric  with  its  pulley,  at  250  R.P.M. 
Where  the  disks  are  the  widest  apart  the  falling  ore  from  the  feed- 
chute  enters  and  is  pinched  and  shattered  as  the  slit  closes,  the  fines 
being  thrown  out  into  the  housing  and  falling  to  a  discharge-spout  beneath. 
The  machine  is  not  suited  to  grinding  sticky  ores,  which  tend  to  adhere  to 
the  disk.  Friable  ores  work  well  in  it.  Another  defect  of  this  crusher  is 
that  "  tramp  iron  "  is  liable  to  break,  or  stall  the  machine.  This  is  over- 
come by  the  use  of  magnets  at  the  conveyor  belt. 

The  Degree  of  Comminution  of  the  ore  is  one  of  the  most  important 
factors,  particularly  in  the  treatment  of  silver  ores.  The  purpose  of  grind- 
ing is  to  free  the  minerals  from  the  inclosing  gangue  and 
to  reduce  the  mineral  particles  to  such  a  size  that  they 
are  readily  dissolved.  Grinding  should  be  carried  on  in 
such  a  manner  as  to  waste  as  little  power  as  possible  in 
grinding  the  worthless  gangue  and  still  fulfill  the  above- 
named  conditions  to  the  fullest  extent  possible,  since  the 
less  the  mineral  is  protected  by  the  gangue  and  the 
greater  the  surface  exposed  to  the  solution,  the  ,more 
rapid  the  rate  of  dissolution. 

It  then  follows,  for  example,  in  the  case  of  certain 
silver  ores,  that  while  fine  grinding  may  not  produce  any 
greater  ultimate  extraction,  yet  in  general  it  will  ma- 
terially reduce  the  time  of  treatment  necessary.  But 
this  advantage  is  not  always  realized  without  the  dis- 
advantage of  greater  consumption  of  cyanide  arising, 
since  finer  grinding  not  only  causes  a  greater  surface  of 
the  minerals  containing  the  precious  metals  to  be  exposed 
to  the  solution,  but  also  a  greater  surface  of  those  minerals,  if  present, 
which  may  act  as  cyanicides.  Careful  correlation  of  the  time  of  treat- 
ment with  degree  of  comminution  may  serve  to  minimize  this  difficulty. 

Selective  grinding  whereby  the  heavy  mineral  particles  are  ground  finer 
than  the  lighter  particles  of  gangue  takes  place  automatically  to  %  greater 
or  less  degree  in  closed  circuits  where  hydraulic  or  mechanical  classifiers 
are  employed  so  that  actually  in  the  majority  of  plants  the  heavy  mineral 
particles  are  ground  finer  than  the  gangue.  The  additional  cost  of  finer 
grinding,  together  with  the  attendant  disadvantages  which  may  arise, 
must  in  each  case  be  carefully  weighed  against  the  additional  extraction 
possible,  or  the  decreased  time  of  treatment  necessary  to  obtain  a  given 
extraction. 

In  grinding  a  sticky  talcose  ore,  the  pebbles  may  become  coated,  the 
noise  dies  down  and  grinding  ceases.  If,  however,  the  feed  is  cut  off  for  a 


FIG.  49. — Disks  of 
Symons  Crusher. 


FINE  GRINDING 


63 


few  minutes  the  pebbles  free  themselves,  the  noise  resumes,  and  grinding 
again  continues. 

Ball  or  Tube-mill  Drive. — Fig.  50  is  a  plan  of  an  excellent  drive, 
because  of  the  use  of  a  flexible  coupling.     It  does  away  with  the  trouble 


Pinion 


75  Horse-power       Chain  Drive 
Motor 

Spur  Geai 
FIG.  50.— Plan  of  Ball-mill  Drive. 

due  to  the  strains  on  the  transmission.  As  arranged  it  drives  a  6-ft.  by 
4J-ft.  ball  mill  at  24  R.P.M.;  a  7  by  12-ft.  tube  or  pebble-mill  at  22  R.P.M. 
Tube-mill  Linings. — To  withstand  the  wear  of  rolls  or  pebbles  on  the 
interior  of  a  ball  or  tube-mill,  special  tough  steel  or  iron  plates  are  pro- 
vided. In  Fig.  51,  A  represents  the  Tonopah  lining  where  longitudinal 


FIG.  51. — Tube-mill  Liners. 


ribs  give  spaces  to  be  filled  with  concrete.  B  is  the  so-called  El  Oro  liner, 
shown  with  its  load  of  pebbles  while  revolving.  The  pebbles  after  a  few 
revolutions  wedge  themselves  into  the  grooves,  affording  a  resistant 
surface  to  the  action  of  the  falling  pebbles.  In  C  we  have  the  Komata 
liner  made  with  heavy  ribs.  These,  during  the  revolution  of  the  mill, 
raise  the  pebbles  quite  high  before  they  can  fall  back. 


64  CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


FINE  GRINDING 

Tube  Mills. — They  are  the  best  of  the  all-sliming  machines,  and  are 
commonly  in  use  in  modern  plants,  either  for  combined  sand  leaching 
and  slime  treatment  in  cyanide  practice,  or  for  all-slime  treatment. 
Fig.  52  gives  a  view  of  such  a  mill.  They  have  been  largely  made  5  ft. 
diameter  by  22  ft.  long,  but  now  mills  6  ft.  diameter  by  14  to  16  ft.  long 
are  preferred.  The  interior  is  lined  throughout  by  thick  steel  plates  which 
last  from  nine  to  twenty-four  months.  The  method  of  the  feed  and  dis- 
charge is  the  same  as  for  the  ball-mills. 

The  speed  varies  from  28  to  34  R.P.M.  The  cylinder  is  filled  half  full 
of  hard  flint  pebbles.  The  falling  and  rolling  of  the  pebbles  on  each  other 
and  upon  the  lining  as  the  cylinder  revolves  grinds  the  material,  in  part  by 
impact,  in  part  by  rubbing  abrasion.  The  pulp,  fed  to  the  mill,  varies 


biG.  52. — Tube-mill. 

in  size  from  J  in.  diameter  to  40-mesh  size,  and  preferably  should  contain 
40  per  cent  water.  In  sliming  to  minus  200-mesh,  the  machine  is  worked 
in  closed  circuit  with  a  classifier,  which  allows  only  the  finer  pulp  to  escape, 
and  returns  the  coarse  material  to  be  reground. 

Dry  Crushing  and  Screening. — Ores  are  dry-crushed  and  screened  in  a 
closed  circuit  as  a  preliminary  to  roasting.  It  is  done  by  means  of  rolls 
and  revolving  screens  or  trommels  in  series,  as  shown  in  the  diagram 
or  flow-sheet,  Fig.  53.  t 

The  ore  supply  from  a  dryer  such  as  the  White-Howell  roaster,  Fig.  73, 
goes  to  the  roughing  rolls  a  which  reduce  it  from  0.75  to  0.25  in.  The 
crushed  ore  is  raised  by  the  elevator  to  a  trommel,  or  separating-screen, 
having  screens  of  |-and  |-in.  aperture,  respectively.  The  first  two-thirds 
of  the  screen  takes  out  all  material  less  than  ^  in.,  and  the  final  size  is  all 
coarser  than  \  in.  We  thus  get  three  products,  an  oversize  from  the  coarser 
screen  which  goes  back  to  the  roughing  rolls  to  be  recrushed,  a  screened 
product  or  undersize,  which  goes  to  the  medium  rolls,  b,  set  at  |  in.  open, 
there  to  be  crushed  and  sent  back  to  the  separating  screen,  and  finally 


STAGE  GRINDING 


65 


an  undersize  through  |-in.  screen,  fine  enough  to  go  to  the  finishing  rolls. 
Until  crushed  so  fine  that  it  passes  the  finest  mesh  the  ore  is  returned  to 
the  trommel.  The  fine  product  of  the  screen  is  raised  by  the  elevator  /' 
and  the  ore  stream  is  equally  divided  between  s"  and  s"',  which  are  pro- 
vided with  30-mesh  wire  cloth.  The  undersize  from  these  trommels  drops 
into  the  storage  bin  m,  while  the  oversize  is  conveyed  to  the  finishing  rolls, 
after  which  it  goes  by  elevator  to  the  finishing  screens.  Thus,  nothing 
enters  the  bin  except  30-mesh,  or  finely  crushed  ore  ready  for  further  treat- 
ment. 

This  system  of  graded  crushing  is  preferable  because  the  final  product 


Finish 


f 

wl 

/^          < 

V 

"|\     Finiahin 

"screens     /p 

Finish!   g  Rolls 


FIG.  53. — Flow-sheet  for  Dry-crushing  or  Screening. 

contains  a  minimum  of  fine  or  slimed  ore  and  being  granular  is  more  easily 
percolated  or  leached.  As  each  piece  or  particle  of  ore  is  crushed  by  a  single 
nip,  the  fine  is  separated  by  the  screen  and  protected  from  unnecessary 
breaking  with  consequent  waste  of  power.  The  chief  costs  in  this  system 
of  dry-crushing  are  those  of  labor,  power,  supplies,  and  repairs.  These 
vary  with  the  tonnage.  The  cost  of  crushing  in  1913  to  30-mesh  size,  in 
preparation  for  roasting  or  leaching  is  50  cents  per  ton,  but  to  this  must 
be  added  overhead  or  general  expense. 


66 


CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


SCREENING 

Screening. — A  mixture  of  coarse  and  fine  mineral,  shaken  through  a 
bar-screen,  a  plate-screen,  or  a  wire-cloth  screen  will  divide  into  two 
pro.ducts,  an  oversize  or  coarse  product,  which  remains  upon  the  screen, 
and  an  undersize  or  minus  product  which  drops  through  its  openings,  and 
which  is  of  the  size  of  the  opening  down  to  the  finest  particles  of  the  ore. 
The  oversize  above  mentioned  may  be  again  crushed  and  fed  upon  the 
screen,  and  by  sufficiently  repeating  the  operation  it  all  eventually  will 
pass  through  as  undersize.  Or  again,  the  undersize  of  a  screen  may  be 
brought  upon  a  finer  one  and  will  then  yield  an  oversize  product,  which  will 
vary  from  the  minus  size  of  the  first  screen  to  the  plus  oversize  of  the 
second.  Thus,  by  using  a  succession  of  screens,  we  obtain  a  series  of 
products,  grades  from  the  coarsest  to  the  finest,  and  a  residual  one  of  all 
finer  than  the  finest  screen. 

Classifying. — Finely  ground  ore  is  called  "  pulp  " — if  dry,  a  "  dry 
pulp,"  if  wet,  a  "  wet  pulp."  It  consists  of  coarse  and  fine  particles 
called  respectively  sand  and  slime.  Such  materials,  when  wet  and  sus- 
pended in  a  liquid,  may  be  separated  into  sand  and  slime  by  means  of 

hydraulic  classifiers,  the  sand 
settling  and  constituting  the 
"  underflow"  or  "spigot" 
discharge  of  such  an  appa- 
ratus, while  the  slime,  re- 
maining in  suspension,  passes 
away  above  at  the  "  over- 
flow." Classifiers  may  be 
of  large  capacity,  and  are 
cheaper  and  more  durable 
than  screens. 

They  may  be  divided  into 
two  kinds,  viz.,  hydraulic 
classifiers  and  mechanical 
classifiers.  i 

Bar  Screen  or  Grizzly. — 
This  is  made  of  bars  6  to  12 
ft.  long,  spaced  about  1J  in. 
apart,  Fig.  54.  Set  at  a  steep 
slope,  as  in  a,  Fig.  134,  it 

receives  the  ore  dumped  upon  it  from  the  mine-car.  The  finer  ore  drops 
through  the  bars  to  the  bin  beneath,  the  oversize  joins  it  after  passing 
through  the  crusher. 

The  Impact  Screen. — For  the  dry  screening  this  screen,   Fig.  55,  is 


FIG.  54. — Grizzlies. 


SCREENING 


67 


well  suited  to  fine  screening.     The  screen  frame,  highly  inclined,  receives 
a  bouncing  or  bumping  motion,  thereby  giving  large  capacity  in  a  small 


The   Impact  Screen 


FIG.  55. — The  Impact  Screen. 


FIQ.  56.— The  Tension  Screen. 

space  with  a  minimum  wear  of  cloth.     The  launder  at  the  head  has  pointed 
cleats  or  guides  which  distribute  the  feed  evenly  over  the  screen. 


68  CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 

The  Revolving  Screen  or  Trommel. — In  Fig.  57  is  a  trommel,  mounted 
on  a  shaft  inclined  commonly  4°  to  5°  to  the  horizontal,  and  revolving 
to  20  times  per  minute.  The  cylindrical  part  is  covered  by  a  punched 
plate  or  by  wire  cloth.  It  has  at  the  left  a  short  receiving  cone  where  the 
ore  enters  a  V-shaped  sheet-steel  housing  or  casing  receiving  the  watery 
undersize,  which  discharges  by  the  spout  A.  A  second  spout  having  a 
dividing  partition  to  prevent  mixing,  discharges  at  B.  Any  portion  of  the 
ore  may  be  said  to  travel  through  the  trommel  in  a  spiral  path.  The 
trommel  is  operated  wet  (sometimes  dry),  wash-water  being  fed  on  the 
outside  of  the  up-coming  side  by  a  spray  pipe.  A  practical  size  for  a  trom- 


FIG.  57. — Revolving-screen*  or  Trommel. 

mel  would  be  3  ft.  diameter  by  6  to  8  ft.  long  with  a  capacity  through  |-in 
holes  of  200  tons  per  day. 

CLASSIFYING 

Classifiers. — These  may  be  divided  into  two  types,  the  hydraulic  and 
the  mechanical  classifiers.  In  the  first  we  include  the  Caldecott  ajid  the 
Allen  cones;  in  the  second  the  Dorr  and  the  Akins  classifiers.  Of  the 
cones,  the  Caldecott  needs  much  attention,  in  order  at  all  times  to  pro- 
duce a  uniform  feed  to  the  tube-mills.  It  returns  much  fine  material  and 
is  a  continuous  source  of  trouble.  The  Allen  cone,  being  automatic  in  'ts 
action,  is  not  nearly  so  open  to  these  objections.  The  Dorr  and  Akins 
classifiers  require  little  attention,  and  will  deliver  a  feed  with  20  to  30  per 
cent  moisture.  The  chief  objection  to  these  machines,  when  dealing  with  a 
heavy  sulphide  concentrate,  is  the  delivery  to  the  tube-mill  of  a  consid- 
erable amount  of  fine  material  that  does  not  require  finer  grinding. 


CLASSIFYING 


The  Caldecott  Diaphragm  Cone-classifier.— This  hydraulic  classifier, 
Fig.  58,  6  ft.  diameter  by  9  ft.  high,  is  popular  in  South  Africa  for  supplying 
a  feed  to  tube-mills.  Its  advantage  over  the  ordinary  cone  type  is  that  it 
has  a  diaphragm  which  sustains  the 
sands  accumulating  above  it  in  the  tank 
nearly  to  the  inlet  spout  as  shown  in  the 
view  of  the  Allen  Cone,  Fig.  59,  while  at 
the  bottom  is  a  cut-off  valve,  that  can  be 
adjusted  to  regulate  the  discharge.  The 
pulp  feed  of  the  launder  passes  through 
a  screen  to  remove  clips  and  floating 
objects,  and  falls  in  the  cylindrical  inlet 
spout  upon  a  deflector  to  gently  settle 
out  the  sands.  The  slime  overflows  over 
the  entire  circumference  of  the  cone  into 
the  peripheral  launder. 

The  Allen  Cone.— This  classifier,  Fig. 
59,   like   the  Caldecott,  is  of  the  same 
settling-basin   type,  not  using   hydraulic 
water,    the   sand    settling   in   the    basin    FIG.  58.— The  Caldecott  Diaphragm- 
while  the  slime  escapes  by  the  extended  cone, 
peripheral  launder.      The  settling  solids, 

forming  a  sand-bed,  are  removed  as  they  accumulate  by  an  automatic 
float-controlled  spigot-valve  29.  The  feed  drops  downward  through  the 
inlet  pipe  or  spout  1,  impinges  on  its  casting,  which  quiets  the  flow,  and 
seeks  exit  through  the  truncated  cone  3.  When  the  sand-bed  accumulates 
to  near  the  mouth  of  the  cone  it  partly  obstructs  the  outward  flow,  causing 
the  watery  pulp  feed  to  rise  in  the  cone  and  carry  with  it  the  float  5. 
By  means  of  the  lever  13  the  link  28  actuates  the  ball  valve  29  and  the 
settled  solids,  even  as  large  as  J-in.  pieces,  begin  to  flow  out  until  the  level 
of  the  solids  in  the  basin  drops,  and  with  it  the  float  5,  again  permitting 
free  flow.  The  superiority  of  this  cone  over  other  hydraulic  classifiers 
is  in  the  automatic  control  of  the  sand  discharge,  since,  when  properly  fed 
and  adjusted  it  needs  no  attention. 

The  Akins  Classifier. — In  this  machine,  Fig.  60,  as  in  the  Dorr  classi- 
fier, which  follows,  the  sand  is  elevated  from  an  inclined  settling  tank  or 
trough  by  mechanical  means,  in  this  case  by  a  spiral  or  screw,  which  fits 
closely  to  the  semicircular  bottom  of  the  trough,  scraping  the  sands 
upward  to  the  discharge  at  the  upper  end  of  the  trough,  while  the  slime  is 
agitated  and  overflows  a  diaphragm  at  the  lower  end  to  the  discharge 
pipe  R. 

The  Dorr  Bowl  Classifier. — This  is  a  Dorr  classifier  to  which  has 
been  added  a  bowl  or  tray,  as  shown  in  Fig.  61,  which  receives  the 


70  CRUSHING,  GRINDING,  SCREENING  AND  CLASSIFYING 


16 


FIG.  59. — The  Allen  Cone. 


FIG.  60.— The  Akins  Classifier) 


CLASSIFYING 


71 


feed  at  its  center,  the  slimed  pulp  overflows  the  periphery  into  the  over- 
flow launder,  while  the  sand  is  plowed  to  the  center  of  the  bowl  and 
discharged  through  a  comparatively  small  opening  into  the  main  inclined 
settling  tank.  Here  it  is  raked  to  the  upper  end  of  the  tank  by  means 
of  two  sets  of  reciprocating  scrapers.  The  sand  settling  to  the  bottom 
is  gradually  advanced  up  it  by  the  forward  motion  of  the  scrapers.  The 
mechanism  then  lifts  these  scrapers  which  drop  to  the  bottom,  then  make 


FIG.  61. — The  Dorr  Bowl-classifier. 

the  next  forward  stroke.  The  coarser  material  (the  sand),  emerging  from 
the  solution  is  discharged  at  the  upper  end  still  containing  25  per  cent  of 
moisture.  Wash  water  or  solution  is  admitted  into  the  main  classifier 
tank  and  flows  underneath  the  bowl  and  up  through  its  central  opening, 
counter-current  to  the  sand.  Usually  about  1  ton  of  back-flow  wash 
per  ton  of  sand  is  sufficient  to  remove  all  slime  from  the  sand.  The  agita- 
tion is  just  enough  to  prevent  settling  of  the  slime  which  now  overflows  at 
the  periphery  of  the  bowl, 


CHAPTER  VI 


METALLURGICAL  FURNACES 

Classification. — These  may  be  divided  into  two  general  types — the 
shaft  furnace .  and  the  reverberatory  furnace.     The  first  consists  of  a  ver- 
tical shaft  or  chimney,  as  in  Fig.  62, 
while  in  the  second  the  charge  is 
kept  separate  from  the  fire. 

THE  SHAFT-FURNACE 

These  are  of  two  kinds,  the 
wind-furnace  using  natural  draft 
as  in  Fig.  64  and  the  blast-furnace 
Fig.  65. 

The  Wind-furnace. — This  is  used 
in  the  laboratory  where  it  is  desired 
to  melt  a  charge  in  crucibles.  When 
converted  for  use  as  a  muffle  furnace, 
the  muffle  is  inserted  at  the  side, 
and  is  surrounded  by  the  burning 
coke,  which  gives  it  a  temperature 
high  enough  for  melting  and  cupel- 
ling. As  a  variation  from  this,  the 
flame  from  a  soft  coal  fire,  carried 
below  the  muffle,  will  sufficiently 
heat  it.  The  wind-furnace  is  used 
in  metallurgical  operations  for  melt- 
ing down  a  product  in  plumbago  or 
clay  crucibles  of  100  Ib.  capacity  or 
over,  these  being  embedded  in  the 
coke.  Indeed  crucible  steel  is  still 
melted  in  this  way. 

The  Blast-furnace.  — Fig.  65 
shows  the  elements  of  a  blast-furnace 


FIG.  62.— Cupola  Blast  Furnace. 


and  the  way  it  is  charged  with  coke  and  materials  of  the  charge  in 
alternate  layers.  It  is  seen  that  the  charge  and  fuel  are  in  intimate  contact 
with  one  another  so  that  not  only  is  the  charge  effectually  heated,  but  is 
brought  within  the  reducing  action  of  the  fuel. 

72 


THE  BLAST-FURNACE  73 

The  Cupola-furnace. — Fig.  62  is  a  cylindrical  blast-furnace  such  as  is 
used  in  foundries  for  the  melting  of  pig-iron  for  castings,  and  also  in  small 
plants  for. the  treatment  of  copper  ores.  The  cylindrical  shaft  is  water- 
jacketed  of  steel  plates.  At  the  bottom  are  seen  the  two  half  drop-doors, 
which,  when  the  furnace  is  running  are  swung  up  in  position  to  close  the 
bottom  of  the  furnace.  On  them  is  laid  the  bottom  lining  of  the  firebrick. 
The  crucible  C,  as  high  as  the  wind  box  E,  is  also  lined  with  4J  in.  of  fire- 
brick. The  shaft  above  this  point,  to  as  high  as  the  feed  floor,  is  double, 
with  a  space  of  4  in.  between  the  inner  and  the  outer  plates.  Since  this 
space  is  filled  with  water  it  is  called  a  "  water-jacket."  The  water  here 
would  soon  reach  the  boiling  point  were  it  not  for  its  circulation  within, 
the  cold  water  entering  both  at  the  top  and  bottom  of  the  jacket  by  the 
supply  pipe  e,  and  the  hot  water  leaving  by  the  overflow  pipe  d.  Thus, 
while  the  inner  shell  is  in  contact  with  the  highly  heated  contents  of  the 
furnace,  it  is  kept  cool  and  is  not  melted  or  attacked  by  the  slag,  which 
indeed  coats  it.  The  lower  part  of  the  water-jacket  is  surrounded  by  the 
wind  box  E,  as  shown.  There  are  six  tuyere-openings  through  the  water- 
jackets  into  the  furnace,  with  openings  opposite  them  in  the  wind-box  as 
shown,  these  latter  closed  by  covers  having  mica-covered  peep-openings 
by  which  the  condition  of  the  tuyeres  can  be  observed.  In  case  of  the 
stoppage  of  a  tuyere,  the  cover  is  removed  and  a  punch  bar  driven  in  to 
remove  the  obstruction.  The  base-plate  rests  upon  four  cast-iron  columns, 
and  sustains  the  entire  structure.  In  the  top  T  is  the  feed-door  by  which 
coke  and  charge  are  put  into  the  furnace,  the  feed-floor  being  at  the  level 
of  the  sill  of  the  feed-door  as  shown  in  Fig.  190.  This  figure  also  shows  how 
the  gases  and  smoke  are  carried  away  by  a  branching  "  downtake  "  to  a 
dust-flue,  or  when  desired  let  go  above  the  roof  of  the  blastfurnace  building. 

Other  types  of  blast-furnaces  are  to  be  found  in  Figs.  152  and  155  for 
iron,  in  Fig.  195  for  copper,  and  in  Figs.  257,  258  and  261  for  lead. 

THE  REVERBERATORY  FURNACE 

This  consists  essentially  of  an  enclosed  fireplace  or  firebox  at  the  right- 
hand  side  of  the  melting  hearth  B,  Fig.  63  or  as  shown  at  a  of  Fig.  71. 
Here  the  fuel  is  burned,  the  products  of  combustion  and  the  flame  being 
drawn  over  the  hearth  to  a  chimney.  The  hearth  is  covered  by  an  arched 
roof  so  as  to  reverberate  or  throw  down  the  heat  upon  the  charge  placed 
on  the  hearth  or  sole  of  the  furnace.  This  space  is  sometimes  called  the 
"  laboratory,"  that  is,  the  enclosure  where  the  labor  or  work  upon  the 
charge  is  performed. 

In  case  the  furnace  is  designed  for  roasting  ore,  then  the  hearth  is  flat 
and  level  with  the  door  sills,  but  when,  as  in  reverberatory  smelting,  the 
contents  of  the  furnace  are  melted,  the  hearth  must  be  dish-shaped  or 
hollowing  and  beneath  the  sill  level  so  that  the  molten  contents  are  retained. 


74  METALLURGICAL  FURNACES 

,  Fig.  63  gives  a  general  idea  of  the  appearance  of  a  reverberatory  roast- 
ing and  melting  furnace,  a  portion  being  broken  away  to  show  the  hearth, 
side  walls,  and  arched  roof.  It  brings  out  clearly  the  way  the  furnace  is 
"  ironed,"  that  is,  tied  together  to  resist  the  thrust  of  the  arch  and  the 
expansion  of  the  brick-work,  as  the  furnace  is  heated.  The  ironing  con- 
sists of  upright  buck-staves  of  railroad  rails,  one  on  each  side  of  each 
door,  tied  across  at  top  and  bottom  (the  lower  tie-rod  beneath  the  floor 
level)  by  IJ-in.  tie-rods.  This  is  also  shown  in  Fig.  71.  Other  buck- 
staves,  set  at  the  ends  of  the  furnace  and  of  the  melting  section  or  fuse-box 
B,  are  tied  to  resist  longitudinal  expansion.  The  furnace  smoke  passes 
away  through  an  outlet  port  at  the  end  of  the  furnace  to  the  flue  F  and 
thence  to  the  stack  common  to  several  furnaces.  The  fuse-box  or  hearth 
at  B  has  two  counterbalanced  lifting  side  doors,  as  has  also  the  fire-box, 
where  is  also  to  be  seen  the  ash-pit  beneath.  The  door  openings  DD  have 


FIG.  63. — Hand-roasting  Reverberatory  Furnace. 

simple  sheet-iron  covers.  Between  them  is  seen  a  flat  plate  with  an  arched 
back  rib  to  stiffen  it.  This  takes  the  thrust  of  the  arch  between  the  doors. 
In  Fig.  213  is  shown  the  heavy  ironing  needed  for  a  large  smelting  rever- 
beratory furnace,  where,  due  to  the  high  melting  heat  needed  the  expansion 
strains  are  great.  The  buck-staves  then  are  6-  and  8-in.  I-beams  with  1J- 
to  2-in.  tie-rods.  One  must  note  here  that  as  the  furnace  heats  up  the  tie- 
rod  nuts  are  slacked  off  a  little,  and  in  cooling  down  they  are  tightened  up — 
all  to  reduce  severe  strains.  The  open-hearth  steel  furnace,  Fig.  176, 
is  similarly  tied. 

In  the  furnace,  Fig.  63,  the  main  hearth  of  the  furnace  is  used  foii  roast- 
ing. This  completed,  the  charge  is  pushed  down  through  a  broad  port 
into  the  fuse  box  B,  there  to  be  sintered  or  melted.  When  melted  it  is 
withdrawn  by  means  of  rabbles  (hoes)  at  the  side  doors  into  wheel- 
barrows set  beneath  the  door-plate. 

The  above  description  refers  to  a  coal-fired  furnace.  There  are  three 
methods  of  firing  as  described  under  head  of  "  Reverberatory  matte  smelt- 
ing," i.e.,  direct  or  coal-fired,  pulverized  coal  and  oil  firing.  To  this  we  may 
add  producer-gas  firing,  used  for  open-hearth  furnaces  in  steel-making. 


CHAPTER  VII 
COMBUSTION 

PRINCIPLES    OF    COMBUSTION 

Combustion,  as  generally  understood,  may  be  defined  as  a  vigorous 
chemical  combination,  attended  with  the  production  of  light  and  heat. 
To  start  combustion,  the  fuel  must  be  (1)  brought  to  the  temperature  of 
ignition;  (2)  it  must  be  maintained  at  this  temperature;  (3)  a  sufficient 
supply  of  air  must  be  provided ;  and  (4)  the  products  of  combustion  must 
be  removed. 

A  jet  of  gas,  burning  as  it  issues  from  a  tube,  begins  to  take  fire  one  or 
more  inches  from  the  tube,  and  continues  to  burn  as  rapidly  as  the  mole- 
cules of  the  gas  come  in  contact  with  those  of  the  air.  In  an  open-hearth 
furnace  a  current  of  heated  gas  and  one  of  air  mingle  gradually,  and  do  not 
become  fully  mixed  and  inflamed  until  a  few  feet  from  the  outlet  ports. 
It  is  the  same  in  a  reverberatory  furnace  especially  where  there  is  but  little 
more  air  than  required  for  perfect  combustion.  The  furnace,  even  if  100 
ft.  long,  may  be  filled  with  flame  from  end  to  end,  showing  that  gas  at  the 
distant  end  is  still  combining  with  air  and  burning.  If,  in  the  fire-box  of 
such  a  furnace,  we  carry  a  thick  coal  fire,  not  less  than  18  to  24  in.  deep, 
the  air  passes  through  the  fire  freely.  The  flame  is  then  shorter,  since 
the  hydro-carbon  gas  is  speedily  consumed.  The  long  flame  is  desirable 
where  we  wish  to  extend  combustion  through  the  furnace  and  not  to  pro- 
duce so  intense  a  heat  near  the  fire.  If  fuel-oil  be  fed  into  an  oil-burner  or 
injector,  where  it  is  broken  up  by  a  jet  of  steam  or  air  into  a  fine  spray, 
and  then  blown  into  a  red-hot  combustion  chamber,  it  will  burn  like  gas 
with  an  intense  heat.  Finely  powdered  coal  projected  in  the  same  way, 
and  with  a  sufficient  supply  of  air  (preferably  preheated)  burns  rapidly, 
resembling  gas,  and  furnishes  abundant  heat  and  high  temperature. 

Thus,  to  promote  rapid  combustion,  the  fuel  must  be  in  such  form  as  to 
afford  plenty  of  contact  to  the  air.  A  piece  of  charcoal  of  large  size  burns 
readily,  because,  being  porous,  the  air  readily  finds  a  way  to  penetrate 
it;  while  a  lump  of  anthracite,  being  dense,  burns  more  slowly.  Paper  and 
kindling  wood  expose  a  large  surface  to  air,  and  hence  ignite  readily  and 
burn  rapidly.  Paper  in  books,  and  fabrics  in  bales,  burn  with  difficulty. 
They  may  pass  through  fire  only  singed  on  the  outside.  Light  lumber  and 
boards  burn  readily,  while  heavy  beams  of  wood  resist  a  fire,  with  super- 

75 


76  COMBUSTION 

ficial  charring.  A  thick  layer  of  sawdust  or  fine  coal,  thrown  on  a  fire 
may  extinguish  it.  Hence,  in  firing  up,  such  fine  materials  should  be  added 
sparingly,  and  used  with  lump  coal  or  pieces  of  wood,  to  make  passages 
or  cavities,  through  which  the  air  may  pass.  Finally,  as  may  be  seen,  a 
common  error  made  in  fire-building  is  that  while  an  abundant  supply  of 
fuel  may  be  present,  insufficient  provision  is  made  for  the  free  passage  of 
air  through  the  fuel.  A  good  draft  and  a  sufficiently  large  exit-flue  must 
be  provided  to  carry  away  the  products  of  combustion. 

Flame  is  gas  undergoing  combustion.  Soft  coal  and  wood  burn  with  a 
flame  because  the  heat  from  burning  distills,  or  drives  out,  the  hydro- 
carbon gas  which  is  formed.  Anthracite,  coke,  or  charcoal  burns  with 
little  or  no  flame;  while  hydrogen  burns  with  a  non-luminous,  though  very 
hot  flame. 

The  temperature  of  ignition  or  of  kindling  varies  according  to  the 
volatile  constituents  of  the  fuel.  Thus  bituminous  coal,  wood,  and 
ordinary  charcoal  will  kindle  at  toward  400°  C.  Anthracite  and  coke 
are  hard  to  start,  kindling  at  700°  C.,  a  full  cherry  heat. 

COMBUSTION  IN  THE  AIR  AND  IN  THE  BLAST-FURNACE 

The  Natural-draft  Furnace. — As  an  illustration  of  what  takes  place  in 
a  deep  fire,  let  us  consider  a  fire  of  glowing  coke,  Fig.  64.  Here  the  air 
enters  through  the  grate-bars,  and,  at  the  first  instant,  in  contact  with  the 
glowing  fuel,  produces  carbon  dioxide, 

C+20  =  C02,.     .    v    .     .     ;     .     .     .     (1) 

with  the  development  of  a  large  amount  of  heat.  Between  2  and  4  in. 
above  the  grate  (with  a  clear  fire)  we  may  expect  to  find  the  highest  tem- 
perature. This  forms  the  zone  1,  Fig.  64,  and  hence  in  a  crucible  furnace 
the  bottom  of  the  crucible  should  be  set  4  in.  above  the  grate  to  get  the 
full  effect  of  the  heat  of  the  fire. 

As  we  go  upward,  the  CO2  in  excess  acts  on  the  glowing  carbon  and 
dissolves  or  combines  with  it  as  follows: 

C02+C  =  2CO.     ''j -   .  ;  Y  -.    .    ;.'  '  I  .     (2) 

This  reaction  is  accompanied  by  the  absorption  of  heat,  and  thus  zone  2, 
Fig.  64,  is  cooler  than  the  one  below.  In  the  zone  3,  no  reaction  takes  place, 
the  fuel  being  simply  heated  by  the  ascending  gases.  A  little  air  may  pass 
along  the  walls,  and  issuing  above  the  surface  of  the  fuel,  and  mixing  with 
the  CO  gas,  burn  a  small  portion  of  it  to  carbon  dioxide  with  a  blue  flame 
thus: 

CO+O  =  CO2 (3) 

This  reaction  also  is  a  heat-producing  one. 


COMBUSTION 


77 


We  finally  get,  with  a  thick  fire,  a  mixture  of  gases  of  a  composition 
much  like  the  following: 

N  70  per  cent,  CO  25  per  cent,  CO2  2.5  per  cent,  O  0.5  per  cent  and 

H  1.0  per  cent. 

The  presence  of  the  hydrogen  is  due  to  the  decomposition  of  the  moisture 
in  the  air.  This  mixture  of  gases  can  be  made  in  a  gas-producer  (see 
Fig.  7),  and  for  that  reason  it  is  called  producer-gas.  Because  of  its 
content  of  CO  it  can  be  burned  according  to  Equation  (3),  and  used  as  a 
fuel  for  any  purpose  of  heating.  To  burn  the  gas  completely,  and  to  get 


FIG.  64. — Wind  Furnace. 


FIG.  65. — Cupola  Furnace. 


the  most  heat  from  it,  the  thickness  of  the  fire  should  not  be  greater  than 
is  shown  in  zone  1,  Fig.  64. 

Combustion  in  the  Cupola  Furnace. — Fig.  65  represents  a  foundry- 
cupola  charged  with  alternate  layers  of  coke  and  pieces  of  pig  iron. 
Here  the  object  is  to  melt  the  iron,  collecting  it  in  a  pool  or 
bath  at  the  bottom  in  the  crucible  of  the  furnace,  shown  in  the 
illustration.  Air  is  forced  into  the  furnace,  and  fills  all  the  voids, 
rising  through  the  charge  chiefly  in  the  passages  or  openings  offering  the 
least  resistance.  In  an  iron  blast-furnace  (see  Fig.  154)  the  rate  of  upward 
velocity  approximates  6  ft.  per  second.  As  air  meets  the  burning  coke, 
combustion  takes  place  according  to  Equation  (1),  producing  a  white  heat. 
This  action,  in  a  cupola  of  36  to  48  in.  diameter,  extends  upward  about 
3  ft.  from  the  tuyeres,  the  upper  limit  of  zone  1,  Fig.  65.  As  the  gases 
enter  the  zone  2,  the  CO2  just  formed  is  decomposed  by  contact  with  the 
hot  coke  and  forms  CO,  according  to  reaction  (2),  the  change  being  nearly 
complete  at  the  upper  limit  of  zone  2.  In  the  upper  zone  no  change  takes 
place  in  the  gases,  and  they  impart  their  heat  to  the  cold  charge,  which  is 


78  COMBUSTION 

being  supplied  as  fast  as  the  ore  sinks  below  the  required  level.  In  this 
cupola,  where  the  operation  is  one  only  of  melting,  to  attain  the  greatest 
economy  of  fuel,  the  coke  should  be  dense,  the  pieces  large,  and  the  blast 
abundant  to  supply  plenty  of  air.  Thus,  burning  the  coke  is  deferred  to  the 
last,  less  CO  is  formed,  and  the  combustion,  performed  largely  in  zone  1, 
is  more  nearly  complete,  developing  the  largest  possible  amount  of  heat. 

From  Equation  (1)  we  find,  that  to  burn  one  pound  of  carbon  to  CO2, 
and  thus  with  the  greatest  development  of  heat,  there  is  needed  2.66  Ib. 
oxygen,  or  11.6  Ib.  air,  since  air  contains  23  per  cent  oxygen  by  weight. 
At  the  sea-level  12.4  cu.  ft.  air  weighs  1  Ib.  This  makes  143.8  cu.  ft.  or, 
in  round  numbers,  150  cu.  ft.  air  per  pound  of  carbon.  Ordinary  coke 
contains  85  per  cent  carbon,  thus  requiring  122  cu.  ft.  air  per  pound  of 
such  coke.  While  in  theory  12  Ib.  air  should  be  sufficient  per  pound  of 
coal,  it  has  been  found  that  excess  is  needed  for  complete  combustion.  For 
natural  draft,  using  a  thin  fire,  18  to  24  Ib.  air  has  given  the  most  satis- 
factory results,  and  where  air  is  forced  into  a  closed  ash-pit,  and  through 
the  fire-bed  (undergrate  blast),  then  16  Ib.  air,  or  even  less,  is  sufficient. 

Figuring  from  Equation  (2)  in  the  same  way,  we  find,  per  pound  of 
carbon  1.33  Ib.  of  oxygen  required,  or  of  air  5.79  Ib.,  equal  to  71.8  cu.  ft. 
Upon  the  basis  of  coke  containing  85  per  cent  combustible  matter,  61  cu. 
ft.  air  is  required  per  pound  of  fuel  when  burned  to  CO. 

Chimneys  or  Stacks. — In  a  furnace  reaction  not  only  is  it  necessary  that 
the  reaction  elements  be  present  in  mutual  contact,  but  that  the  products 
of  the  reaction  be  removed  as  fast  as  formed.  Under  conditions  other  than 
this  the  reaction  ceases.  A  draft,  therefore,  must  be  provided,  to  carry 
away  the  waste  gas,  and  to  expel  it  into  the  atmosphere.  This  draft  may 
be  natural  or  forced.  To  insure  obtaining  a  sufficient  draft  in  a  chimney, 
the  gases  must  be  delivered  into  the  stack  while  hot.  A  temperature  of 
200°  C.  is  ample  for  this,  but  since  the  work  of  most  furnaces  is  done  at  a 
temperature  higher  than  a  red-heat,  the  excess  may  be  utilized  to  generate 
steam  by  conducting  the  gases  through  waste-heat  boilers  before  entering 
the  stack. 

In  a  reverberatory  furnace,  used  for  smelting,  the  quantity  and  intensity 
of  the  heat  depend  upon  the  amount  of  coal  burned  per  hour.  TFIJLS  varies 
between  18  and  40  Ib.  per  square  foot  of  grate  area  and,  to  burn  it  com- 
pletely, there  will  be  needed  150  cu.  ft.  air  per  pound  of  coal  consumed. 
In  these  furnaces  the  gases  escape  at  temperatures  between  300°  and  1 100° 
C.,  and  move  with  a  velocity  of  12  to  20  ft.  per  second.  For  illustration, 
take  the  furnace  Fig.  122,  with  a  grate-area  of  112  sq.  ft.,  a  consumption 
of  30  Ib.  of  coal  per  square  foot  of  grate-area  per  hour,  needing  514,000 
cu.  ft.  of  free  air.  Allowing  a  temperature  in  this  instance  of  1000°  C., 
and  a  draft  velocity  of  20  ft.  per  second  at  this  high  temperature,  and 
knowing  that  these  gases  expand  7273  of  their  volume  for  each  degree 


CHIMNEY  DRAFT  79 

above  0°  C.,  we  find  the  volume  at  1000°  C.  to  be  10°^j"273=4.7  times 

27o 

the  volume  at  0°  C.  Assuming  the  temperature  of  the  outside  air  to  be 
0°  C.,  we  shall  have  as  the  volume  of  hot  gas  per  hour  2,368,800  cu.  ft. 
At  20  ft.  per  second,  or  72,000  ft.  per  hour,  this  will  be  an  area  of  stack  of 

9  OCQ  o(\f\ 

72000    =  3°  Sq'  ft'     The  aCtUal  area  iS  32  Sq*  ft' 

The  total  pull,  or  suction,  that  a  chimney  can  produce,  assuming  it 
to  be  filled  with  hot  gases,  is  due  simply  to  the  ascensive  force  of  the 
gas  measured  by  the  difference  between  its  weight  and  the  weight  of  an 
equal  volume  of  the  cold  air  outside.  To  maintain  the  velocity  of  the 
gas  in  the  stack,  it  has  been  found  that  a  suction,  or  "  pull/'  of  0.4  to 
0.8  in.  of  water,  as  measured  by  a  water-gauge,  is  needed.  Taking  a  draft 
of  0.6  in.  in  the  above  instance  and  adding  0.1  in.  for  friction  in  the  chim- 

62  5x0  7 
ney,  we  have  0.7  in.  water  equal  to  -  ^—- — —  =  3.647  Ib.  per  sq.  ft.     A 

cubic  foot  of  air  at  0°  C.  weighs  0.0807  Ib.  (12.4  cu.  ft.  per  Ib.).  The  gas 
inside  the  stack  has  a  specific  gravity  of  1.03  weight  of  air  being  unity, 

thus  making  the  weight  when  heated  -  '  '  —  =  0.0177  Ib.  per  cu.  ft. 
Hence  we  have  the  difference  (0.0807-0.0177  =  0.063)  as  the  ascensive 

O  R.A  *7 

force  per  foot  of  height.     The  stack  should  therefore  be    '  ^0=58  ft.  or 

0.063 

approximately  60  ft.  high. 

From  the  above  calculation  it  appears  that  the  draft-pressure  varies 
with  the  temperature  and  height  of  the  chimney.  The  velocity  of  the  gas, 
or  the  amount  of  air  passing  through  the  fire  per  hour  at  a  given  tempera- 
ture, varies  as  the  square  root  of  the  height  of  the  stack,  in  accordance 
with  the  equation  V  =  ^/2gh;  g  being  acceleration  due  to  gravity  and  h 
head  in  feet.  Thus  a  stack  100  ft.  high  would  increase  the  velocity  only 
1.31  times  more  than  our  58-ft.  stack  calculated  above.  The  volume  of 
gas  increases  directly  with  the  temperature,  while  the  velocity  varies  with 
the  square  root  of  this,  hence  there  is  a  point  of  maximum  discharge,  at 
273°  C.  At  a  higher  temperature,  while  the  velocity  increases,  the  weight 
of  the  gas  on  the  contrary  diminishes. 


TEMPERATURE    OF    COMBUSTION 

By  this  is  meant  the  temperature  of  gases  resulting  from  combustion 
under  ordinary  atmospheric  pressure.  We  can  calculate  this  when  we 
know  the  calorific  power  of  the  fuel,  and  the  total  weight  and  mean  specific 
heat  of  the  resultant  gases.  The  specific  heat  between  0°  C.  and  the 
temperature  of  combustion  increases  as  is  shown  in  the  diagram,  Fig.  68. 


80 


COMBUSTION 


Flame  Temperature. — As  an  example  of  the  use  of- the  following  table/ 
let  us  find  the  maximum  temperature  of  combustion  obtained  in  burning 
one  pound  of  coke  of  85  per  cent  carbon,  using  the  theoretical  amount  of 
air  or  9.86  Ib.  (since  pure  carbon  requires  11.6  lb.),  neglecting  the  loss  of 
heat  in  the  adjoining  walls  of  the  furnace.  Since  by  weight  there  is  77 


FIG.  67. — Mahler  Bomb  Calorimeter. 


0.1 


0.3 


0.5 


0.7 


0.9 


2400 

2200 

2000° 

1800° 

1600° 

140d 

1200° 

1000° 

800° 

600° 

400C 

200° 


0.0 


0.2  0.4  0.6  0.8 

FIG.  68. — Specific  Heat  of  Gases. 


1.0 


per  cent  nitrogen  in  air,  there  will  be,  in  the  mixed  gas  resulting  from  com- 
bustion, 7.6  lb.  nitrogen,  and  3.11  lb.  CO2  (0.85  lb.  carbon,  and 
ff  X0.85  =  2.26  lb.  oxygen).  We  assume  a  temperature  of  2000°  C.,  which 
is  as  near  the  desired  one  as  we  can  judge.  From  Fig.  68  we  find  for  2000° 
the  specific  heat  to  be,  for  N2,  0.281  and  for  C02,  0.364.  We  then  have: 


FLAME  TEMPERATURE  81 

7.6  Ib.  nitrogen @  0.281  =  2. 133 

3.11  Ib.  carbon  dioxide @  0.364=  1 . 132 


3.265 

The  number  of  calories  necessary  to  raise  the  entire  gaseous  product 
from  1  Ib.  of  coke  one  degree  will  then  be  3.265.  The  heat  developed  by 
the  burning  of  the  coke,  being  6800  cal.,  the  temperature  of  combustion  is 

f\Q(\f\ 

=  2080°  C.     The  specific  heat  of  the  gases  at  this  temperature  being 


so  nearly  the  same  as  at  2000°  C.,  we  can  use  it  in  this  calculation.  Were 
the  difference  great,  we  should  have  to  use  the  specific  heat  at  the  exact 
temperature  and  calculate  again  upon  that  basis. 

To  proceed  further,  let  us  find  the  temperature  of  combustion,  or 
flame-temperature  of  carbon  monoxide  burning  to  carbon  dioxide: 

CO+O  =CO2  =  68,000  cal. 
28+16=44 

One  pound  of  CO  will  produce  2440  cal.,  and  will  take  0.572  Ib.  oxygen 

or  -      —=2.48  Ib.  air  containing  1.91  Ib.  nitrogen.     There  is  also  1.57 
0.23 

Ib.  CO2  produced.     Taking  the  specific  heats  from  the  table,  Fig.  4. 

1.91  Ib.  nitrogen  ...................   @  0  .  287  =  0  .  548  cal. 

1.57  Ib.  carbon  dioxide  .........         0.364  =  0.572    " 


1.120    ' 

2440 

for  each  degree  of  rise  in  temperature.     Hence  -    —  =  2200°  C.  =  the  tem- 

l.U 

perature  of  combustion.  Again,  calculating  with  the  increased  specific 
heat  of  the  newly  found  temperature  (2200°),  we  find  2112  to  be  the  exac*, 
number  of  calories,  nearly  identical  with  that  already  found. 

Fuel  Combustion. — In  writing  equations  the  molecular  weight  is  under- 
stood as  in  ordinary  chemical  equations. 

(1)  C+O2  =  97,000 
may  also  be  written 

(2)  C,O2  =  97,000 

with  a  comma  to  indicate  that  the  different  molecules,  so  separated,  unite 
to  form  CO2. 


82  COMBUSTION 

If  oxygen  burns  in  presence  of  an  excess  of  highly  heated  carbon,  then 
carbon  monoxide  is  formed,  and  this  may  be  written 

(3)  C+0  =  CO, 

29,000 

which  indicates  that  by  the  combination  of  the  solid  carbon  with  the 
gaseous  oxygen  29,000  calories  have  been  formed.  Likewise  this  may  be 
written 

(4)  C,O  =  29,000 

Since  the  12-lb.  carbon  gives  29,000  calories,  we  have,  as  the  heat  evolved 
by  the  burning  of  1  Ib.  of  carbon,  2440  calories.  If  we  burn  the  CO  thus 
formed  with  a  sufficient  amount  of  air,  we  have 


(5) 

68,000 

or,  as  also  written,  00,0  =  68,000  calories.     Were  this  written  C,O2  then 
we  would  have  97,000  calories  (Equation  (2)). 
Equation  (5)  may  again  be  written 


(6)  OO+0  =  CO2. 

29,000       97,000  =  68,000 

This  means  that  before  this  reaction  can  take  place,  the  CO  must  be  broken 
up  according  to  the  reaction, 

(7)  CO  =  C+0, 

29,000 

or  again 

(8)  C,O  =  C+O  =  29,000, 

the  minus  sign  meaning  that  in  this  reaction  as  much  heat  has  been  absorbed 
in  the  breaking  up  as  was  earlier  evolved  in  Equation  (3),  in  which  these 
elements  united. 

In  Equation  (6)  the  CO  having  been  decomposed  into  C  and  0,  they 
are  forced  to  unite  with  the  O  to  form  CO2,  evolving  97,000  calories*  The 
net  result  or  the  algebraic  sum  of  the  two  reactions  is  thus  68,000  calories 
as  given  in  Equation  (6). 

Tempera  ures  of  Combustion.  —  The  temperature  of  an  incandescent 
body  may  be  judged  by  the  eye  or  by  an  optical  pyrometer  according 
to  the  color  scale  herewith; 

Lowest  visible  red  ..........  •,.,."  ...................     470°  C. 

Dull  red  ....................................  550  to  625° 

Cherry  red  ......................................      700° 

Light  cherry  red  .................................     850° 


FLAME  TEMPERATURE 

Orange 900°  C. 

Yellow 950  to  1000° 

Light  yellow 1050° 

White 1150° 

Dazzling  white .  .  ....  1500  to  1600° 

These  colors  apply  both  to  flame  and  to  a  heated  body. 


83 


j 


FIG.  69. — Wall  Type  Indicating  Pyrometer. 

Indicating  Pyrometer. — Fig.  69  is  a  Le  Chatelier  pyrometer  com- 
posed of  a  thermo-couple  of  a  platinum  and  a  platinum  rhodium  wire 
placed  at  the  end  of  a  protective  tube  of  porcelain  which  is  thrust  into  the 
fire  or  against  the  heated  object.  The  two  wires  are  continued  as  leads  to 
the  terminals  of  a  galvanometer  graduated  to  indicate  the  temperature. 


CHAPTER  VIII 
METALLURGICAL  THERMO-CHEMISTRY 

.       METHODS  OF  DETERMINING  THERMIC-VALUES 

Definition  of  Metallurgical  Thermo-Chemistry. — This  pertains  to  ques- 
tions relating  to  the  heat  evolved  or  absorbed  when  elementary  substances 
or  their  compounds  combine  in  metallurgical  operations.  By  having  an 
intimate  knowledge  of  these  reactions  we  can  utilize  fuel  and  control  roast- 
ing, smelting,  or  converting  operations  to  the  best  advantage. 

Thus  when  cupelling  rich  lead  in  an  English  cupelling  furnace,  Fig. 
181  A,  the  air  passing  over  the  hot  molten  bath  is  seen  to  aid  in  main- 
taining its  molten  condition,  energetically  uniting  with  the  lead  to  form 
litharge,  and  so  evolving  heat. 

When  wood  or  other  ftiel  is^kindled,  the  properly  supplied  and  regulated 
draft  of  air  maintains  the  combustion  "of  the  fuel  with  evolution  of  much 
heat. 

A  mixture  of  quartz-bearing  ore,  with  fluxes  in  suitable  proportion,  the 
whole  brought  to  a  white  heat,  will  melt  together;  the  heat  being  intensified 
in  so  doing. 

On  the  other  hand,  when  steam  is  passed  through  a  glowing  coke  fire, 
it  is  decomposed,  forming  hydrogen  and  carbon  monoxide,  but  absorbing 
heat  and  cooling  the  fire. 

Units  of  Measurement. — The  amount  of  heat  generated  as  the  result 
of  the  combination  of  elementary  substances  is  given  in  Table  I.     In 
Table  II  is  given  the  amount  of  heat  evolved  as  the  result  of  the  combina 
tion  of  certain  bases  (oxides  of  the  metals),  as  given  in  Table  I  with  silica. 

The  unit  of  measurement  used  is  the  heat  required  to  raise  a  unit  weight 
of  water  one  degree.  Thus,  one  gram  of  water  raised  in  temperature  one 
degree  Centigrade  is  called  a  small  calorie  (cal.)  or  gram-calorie.  One 
kilogram  of  water  raised  1°  C.  is  called  a  large  calorie  (Cal.)  or  kilogram- 
calorie.  One  pound  of  water  raised  1°  Fahrenheit  is  called  a  British 
thermal  unit  (B.t.u.)  and  it  is  but  0.252  of  the  kilogram-calorie.  The 
gram-calorie  is  used  in  small-scale  laboratory  operations.  For  ordinary 
work  the  large  calorie  is  preferred. 

In  this  book  and  in  the  calculations  which  follow  we  shall  use  the 
pound-calorie,  this  being  the  heat  required  to  raise  1  Ib.  of  water  1°  C. 
When  we  use  the  word  "  calorie  "  the  pound  calorie  will  be  understood. 
The  kilogram-calorie  is  2.2  times  greater  than  the  pound-calorie. 

General  Principles. — 1.  The  amount  of  heat  needed  to  decompose  a 

84 


CALORIMETRY  85 

compound  into  its  constituents  is  equal  to  that  evolved  when  that  com- 
pound is  formed  from  those  constituents.  When  a  reaction  takes  place  by 
which  heat  is  absorbed,  as  in  Equation  (8),  it  is  called  "  endothermic." 
On  the  other  hand,  when  heat  is  evolved  in  a  reaction,  as  hi  Equations  (1), 
(3),  and  (5),  it  is  said  to  be  "  exothermic." 

2.  The  heat  evolved  hi  a  chemical  process  is  the  same,  whether  it 
takes  place  directly  or  in  several  steps.     Thus  in  Equation  (3)  the  carbon 
is  burned  to  CO  with  the  evolution  of  29,000  calories.     The  CO  thus  formed, 
when  burned  with  additional  oxygen,  as  in  Equation  (5),  gives  68,000  cal- 
ories, and  the  sum  of  these  two  is  97,000  cal.,  the  same  as  if  the  carbon  had 
been  burned  to  CO2  as  per  Equation  (1) 

In  comparing  reactions  (1)  and  (3),  it  may  be  said  that  in  presence  of  an 
excess  of  oxygen,  reaction  (1)  would  take  place  rather  than  reaction  (3). 
This  is  in  accordance  with  the  law  of  Berthelot,  namely : 

3.  Every  reaction  which  takes  place  independently  of  the  addition 
of  energy  from  without  the   system,  tends  to  form  the  combination 
which  is  accompanied  by  the  greatest  evolution  of  heat. 

Calorimetry. — To  determine  accurately  the  heats  of  combustion  of 
fuels,  or  the  heat  of  formation  of  compounds,  the  Mahler  bomb-calorim- 
eter, Fig.  67,  is  much  used.  It  consists  of  a  steel  shell  or  bomb,  marked 
B,  shown  also  on  an  enlarged  scale  at  the  right-hand  upper  corner  of  the 
illustration.  The  bomb,  holding  about  a  pint  and  weighing  9  lb.,  is  shown 
to  be  closed  by  a  screw-cap,  having  a  stop-cock  threaded  connection  x 
by  which  it  may  be  connected  by  a  flexible  pipe  to  a  cylinder  0,  which  con- 
tains compressed  oxygen  gas.  Within  the  bomb  is  suspended  a  capsule  c 
in  which  is  placed  a  gram  of  the  substance  to  be  tested.  The  cap  is  then 
tightly  screwed  on,  and  oxygen  gas  under  a  pressure  of  300  lb.  per  square 
inch  is  allowed  to  enter.  The  shell  is  next  placed  in  the  calorimeter  D, 
which  contains  a  known  weight  of  water.  The  thermometer  T  is  set  in 
place,  and  the  stirrer  or  agitator  S  is  set  in  motion  to  bring  the  whole 
apparatus  to  the  same  temperature.  The  calorimeter  D  is  placed  within 
a  larger  vessel  A  covered  with  a  thick  layer  of  felt  and  provided  with  a 
thermometer  (not  shown).  The  vessel  A  serves  also  to  support  the 
bracket  G  from  which  the  stirrer  is  suspended.  The  temperature  of  the 
calorimeter  having  been  noted,  the  charge  is  ignited  by  a  coil  of  the  plat- 
inum wire  F.  The  resultant  rise  in  temperature  is  noted  by  the  ther- 
mometer T7.  The  total  heat  developed,  with  certain  corrections,  is  cal- 
culated from  the  weight  of  the  calorimeter  water,  and  from  the  rise  in 
temperature.  In  those  cases  where  the  heats  of  formation  of  oxides  or  of 
silicates  are  desired,  the  net  result  is  accomplished  by  respectively  oxidizing 
or  melting  them  in  the  bomb  with  a  known  weight  of  a  well-determined  fuel. 
The  number  of  calories  evolved  is  the  algebraic  sum  of  those  of  the  desired 
reaction  and  that  of  the  fuel. 


86 


METALLURGICAL  THERMO-CHEMISTRY 


HEATS    OF   FORMATION    OF   THE   ELEMENTS 

Following  are  the  molecular  weights  and  the  heats  of  formation  of  some 
of  the  better-known  chemical  compounds.  From  these  may  be  estimated 
the  heat  developed  in  various  reactions : 

TABLE  I 
HEAT  OF  FORMATION  OF  CHEMICAL  ELEMENTS 

Carbon,  Hydrogen  and  Sulphur 


Formula. 

Molecular  Weights. 

Heat  of 
Formation        Formula. 

Heat  of 
Molecular  Weights.          Formation 

(Calories). 

(Calories). 

C,O 

12  + 

16  = 

28 

29,000 

S,O2 

32  + 

32  = 

64 

69,260 

C,O2 

12  + 

32  = 

44 

97,000 

S,03 

32  + 

48  = 

80 

91,900 

H2,0 

2  + 

16  = 

18 

58,060 

Silicon  and  Phosphorus 

Si,02 

28  + 

32=  60 

196,000 

P205 

68  + 

80  = 

142 

365,300 

Oxides  of  the  Metals 

MgO 

24  + 

16  = 

40 

143,000 

Fe2,03 

112  + 

48  = 

160 

195,600 

Ba,O 

137  + 

16  = 

153 

133,400 

Fe3,O4 

168  + 

64  = 

232 

270,800 

Ca,0 

40+ 

16  = 

56 

131.500 

Sb2,O3 

240  + 

48  = 

288 

166,900 

A12,03 

54+ 

48  = 

102 

392^600 

Sb2,05 

240  + 

80  =  320 

231,200 

Na2,O 

46  + 

16  = 

62 

100,900 

As2,O3 

150  + 

48  = 

198 

156,400 

K2,0 

78+ 

16  = 

94 

98,200 

As2,O5 

150  + 

80  = 

230 

219,400 

Mn,O 

55  + 

16  = 

71 

90,900 

Cu,O 

64  + 

16  = 

80 

37,700 

Mn,O2 

55+ 

32  = 

87 

125,300 

Cu2,O 

128  + 

16  = 

144 

43,800 

Zn,0 

65+ 

16  = 

81 

84,800 

Pb,O 

207  + 

16  = 

223 

50,800 

Fe,O 

56+ 

16  = 

72 

65,700 

Pb,02 

207  + 

32  =  239 

63,400 

Sulphides  of  the  Metals 

Ba,S 

137  + 

32  = 

169 

102,900 

Cu2,S 

128  + 

32  = 

160 

20,300 

Ca,S 

40+ 

32  = 

72 

94,300 

Cu,S 

64  + 

32  = 

96 

10,100 

Zn,S 

65  + 

22  = 

97 

43,000 

Pb,S 

207  + 

32  = 

239 

20,200 

Fe,S 

56+ 

32  = 

88 

24,000 

Sb2,S3 

240+ 

96  = 

336 

34,400 

Carbonates  of  the  Metals 

Ba,C,O3 

137  + 

12  + 

48  = 

197  286,300 

Mg,C,O3 

24  + 

12  + 

48=  84 

269,900 

Ca,C,O3 

40  + 

12+ 

48  = 

100273,800 

1 

Sulphates  of  the  Metals  in  Dilute 

Solution 

K2,S,04 

78  + 

32  + 

64  = 

174  337,700 

Zn,S,O4 

65  + 

32  + 

64  =  161 

248,000 

Na2,S,O4 

46+ 

32  + 

64  = 

142  328,500 

Fe,S,04 

56+ 

32  + 

64  =  152 

234,900 

Ca,S,O4 

40+ 

32  + 

64  = 

136321,800 

Fe2,S3,02 

122  + 

96+192=400 

650,500 

Mg,S,04 

24  + 

32  + 

64  = 

120321,100 

H2,S,04 

2  + 

32  + 

64=  98 

210,200 

A12,S3,02 

56+ 

56+192  = 

342879,700 

Cu,S,O4 

64  + 

32  + 

64  =  160 

197,500 

Mn,S,O4 

55  + 

32  + 

64  = 

151  263,200 

XHg,Au 


Amalgams  of  Gold  and  Silver 
Hg  in  excess  197  2,580     |     XHg,Ag      Hg  in  excess  108          2,470 


HEATS  OF  FORMATION 


87 


TABLE   II 

Heats  of  Formation  of  the  Silicates 


Formula. 

Molecular  Weight. 

Formation  Heat. 

Heat  per  Ib. 
of  Slag. 

FeO,SiO, 

72+  60=   132 

10,600 

80 

2FeO,SiO, 

144+  60  =  204 

22,236 

109 

MnO,SiO» 

71+  60  =  131 

5,400 

41 

BaO,SiO, 

153+  60  =  213 

14,700 

69 

CaO,SiO, 

56+  60  =  116 

17,850 

159 

2CaO,SiO» 

112+  60  =  172 

28,300 

165 

3CaO,SiOi 

168+  60  =  228 

28,250 

125 

SrO2,SiO2 

120+  60  =  180 

17,900 

110 

Al2O3,2SiO2 

102  +  120  =  222 

14,900 

67 

3CaO,Al2O3,2SiOs 

168+102+120  =  390 

33,500 

86 

2H2O,Al2O3,2SiO2 

34+102  +  120  =  256 

43,800 

170 

Li20,Si02 

30+  60=  90 

65,100 

720 

Na2O,SiO, 

62+  60  =  122 

45,200 

370 

CaO,Al2O3 

56+102  =  158 

450 

3 

2CuO,Al2O3 

112  +  102  =  214 

3,300  " 

15 

3CaO,Al2O3 

68+102  =  170 

2,950 

11 

SiO2  35.5  +  FeO  39.7+ 

MnO    1.0+CaO  11.4+ 

MgO   2.7+Al2O3  9.2+ 

CuO    0.4+S  0.4% 

(a  compound  slag)  

133 

FeO57.6;  CaO  12.0;  SiO230.4% 

140 

FeO  40.3:  CaO  28.0:  SiO31.7% 

193 

The  heat  of  formation  of  silicates,  if  we  were  to  start  from  the  elements, 
as  in  Table  I,  commonly  amounts  to  from  2000  to  4000  calories  per  pound 
of  the  compound  thus  formed;  but,  when  the  metals  and  silicon  have 
become  oxidized,  as  commonly  occurring,  most  of  the  heat  of  formation 
has  developed.  If  then  such  oxides  and  silica  are  brought  to  melting 
temperature  they  combine  with  a  farther  development  of  heat,  as  given  in 
Table  II,  which  varies  from  almost  nothing  up  to  720  calories  per  pound 
of  the  silicate  formed,  or  an  average  of  145  calories.  Thus,  in  Table  I, 
the  heat  of  formation  of  FeO  is  given  at  65,700  calories  and  of  silica  at 
180,000,  or  together  245,700  calories,  being  1100  per  pound.  But  when 
the  slag  FeO,  SiC>2  is  formed  only  10,600  calories  is  developed,  equal  to  80 
calories  per  pound  of  the  slag. 

HEAT  EVOLVED  AS  THE  RESULT  OF  ROASTING 

In  the  reaction  C+ 62  =  CO2  (see  Table  I,  Carbon,  Hydrogen,  and 
Sulphur)  one  equivalent,  or  12  Ib.  of  carbon  is  completely  burned  by  com- 
bining with  32  Ib.  of  oxygen,  forming  44  Ib.  of  carbon-dioxide  and  evolving 
97,000  calories.  Dividing  this  by  12  we  have  8080  calories  as  the  result  of 
the  burning  of  1  Ib.  of  carbon. 


CHAPTER  IX 
ROASTING 

By  roasting  we  mean  the  preliminary  treatment  of  ores  by  fire  at 
temperatures  below  their  melting  point  in  order  to  improve  their  condition 
for  subsequent  reduction  or  extraction.  Because  of  the  expense  the  opera- 
tion is  avoided  where  possible. 

We  may  classify  these  operations  into:  (1)  Calcination  or  kiln  roasting; 
(2)  oxidizing  roasting;  (3)  chloridizing  roasting:  (4)  sulphatizing  roasting; 
(5)  sinter  roasting. 

Calcination  or  Kiln-roasting. — Carbonate  ores,  as  of  iron  or  zinc,  are 
charged  in  lump  form  together  with  some  coal  or  wood  into  a  vertical 
kiln.  The  heat  expels  the  contained  moisture  and  carbon  dioxide  of  the 
ore.  In  the  case  of  some  Mesabi  iron  ore,  containing  12  to '15  per  cent  of 
moisture,  freight  is  saved  by  kiln  drying  and  the  ore  is  made  more  porous 
and  accessible  to  the  action  of  reducing  gases  in  the  subsequent  smelting. 
The  grade  of  the  ore  is  raised  and  a  better  price  is  obtained  for  it;  if  a 
siderite  or  iron  carbonate,  the  weight  is  still  further  reduced. 

Oxidizing  Roasting. — This  is  for  the  purpose  of  expelling  the  moisture, 
of  burning  off  the  sulphur,  and  in  the  case  of  certain  refractory  ores,  of 
removing  their  contained  arsenic  or  tellurium.  Thus,  these  are  freed  from 
impurities  which  would  interfere  in  subsequent  smelting  or  leaching.  The 
ore  is  heated  to  burning  temperature  with  free  access  of  air,  and  gives 
a  porous  product,  easily  penetrated  by  reducing  gases  or  by  solutions. 
As  a  treatment  for  leaching,  roasting  destroys  colloids,  so  that  the  roasted 
ore  is  more  easily  leached.  The  product  of  an  oxidizing  roast  is  often 
called  calcines,  though  this  is  better  applied  to  the  product  of  calcination. 

Chloridizing  Roasting. — It  is  performed  in  a  reverberatory  fiirnace, 
common  salt  being  added  at  a  certain  stage  of  the  roasting  to  change  the 
ore  to  an  easily  leached  chloride.  Advantage  is  taken  here  of  the  principle 
of  mass-action,  whereby  the  nascent  chlorine  reacts  on  the  oxidized  metal- 
liferous products. 

Sulphatizing  Roasting. — This  is  a  partial  oxidizing  roast,  some  of  the 
sulphur  being  expelled,  and  the  remainder  changed  to  SOs.  Thus  a  sul- 
phate of  the  metal  is  formed,  which  is  soluble  and  therefore  can  be  extracted 
by  water-leaching. 

Sinter  Roasting. — Sulphide  ore  spread  out  on  a  traveling  grate  is  heated 

88 


CHEMISTRY  OF  ROASTING  89 

by  a  flame  to  burning  temperature.  Air  is  then  drawn  through  it  by  aid 
of  a  suction  fan,  resulting  in  a  vigorous  oxidation,  expulsion  of  the  sulphur, 
and  a  partial  sintering  together  of  the  'ore  particles.  Not  only  is  the  ore 
well  roasted,  but  the  dust  loss  in  the  subsequent  blast-furnace  smelting  is 
greatly  lessened.  Iron  ore  in  too  fine  condition  for  blast-furnace  smelting 
has  been  mixed  with  6  to  10  per  cent  of  its  weight  of  fine  coal  to  sinter  or 
agglomerate  it.  In  the  past  much  fine  ore  has  not  been  so  treated,  but 
has  been  fed  in  the  crude  state  so  that  the  blast-furnace  has  made  a  high 
flue-dust  loss. 

CHEMISTRY  OF  ROASTING 

Chemistry  of  Oxidizing  Roasting. — To  do  good  roasting,  we  should  have 
(1)  heat  sufficient  to  start  the  burning;  (2)  a  final  heat  sufficient  to  drive 
off  the  last  portion  of  the  sulphur;  (3)  preferably  46  Ibs.  of  air  per  pound  of 
sulphur,  (4)  an  extensive  surface  exposed  to  the  air;  (5)  frequent  stirring  in 
order  to  present  to  the  air  fresh  surfaces  for  roasting. 

Let  us  take  the  case  of  an  ore  with  a  silicious  gangue,  containing  the 
sulphides,  pyrite,  chalcopyrite,  blende,  and  galena.  This  is  dropped  upon 
the  hearth  at  the  hopper  end  of  the  furnace,  Fig.  63,  and  then  spread  out. 
Here  the  temperature,  350°  C.  is  sufficient  to  expel  moisture  and  start  the 
reaction  of  combustion.  In  ten  to  fifteen  minutes  burning  begins,  as 
evinced  by  a  blue  flickering  sulphur  flame  that  plays  over  the  surface  of  the 
charge.  The  pyrite  is  thus  decomposed. 

(1)  FeS2+O2+heat  =  FeS+SO2. 

The  first,  loosely  held  equivalent  of  sulphur  is  easily  expelled  and  unites 
with  the  air,  with  the  evolution  of  3220  pound-calories  per  pound  of  sul- 
phur burned.  The  FeS  now  remaining,  together  with  the  other  sulphides, 
begins  to  oxidize.  The  FeS  and  CuS  is  most  easily  oxidized,  while  the  ZnS 
and  PbS  are  the  slowest  in  parting  with  their  sulphur.  Beginning  then 
with  the  FeS  we  have 

(2)  FeS +  30=  FeO  +  SO2, 

23,800  66,400    7 1,000= +113,600 

or  in  words,  the  iron  sulphide  becomes  oxidized  to  ferrous  oxide  with  the 
formation  of  SO2,  the  reaction  being  exothermic,  and  yielding  113, 600 -r- 
32  =  3550  pound  calories  per  pound  of  sulphur  burned.  Cupric  sulphide  of 
the  chalcopyrite  reacts  according  to  the  formula: 

(3)  CuS  +  3O   =   CuO  +  SO2, 

10,200  37,200     71,000= +98,000 

or  per  pound  of  sulphur,  98,000+32  =  3030  Cal.  The  blende  under  the 
action  of  air  and  heat  is  affected  in  the  same  way. 


90  ROASTING 

(4)  ZnO  +  3O  =  ZnO  +  SO2, 

43,000  86,400     7 1 ,000  =  + 1 14,400 

or  per  pound  of  sulphur  present,  3420  Cal.  Galena  roasts  according  to 
the  reaction: 

(5)  PbS  +  30  -   PbO    -f     SO2, 

17,800  51,000        7 1,000  =+104,200 

which  gives  off  3250  Cal.  per  pound  of  sulphur. 

It  will  be  noticed  that  the  heat  evolved  per  pound  of  sulphur  is  much 
the  same  in  each  case,  and  hence  the  sulphide  highest  in  sulphur  yields 
the  most  heat.  These  reactions,  especially  of  blende  and  galena,  are  grad- 
ual during  the  roasting  period.  The  air  acts  chiefly  on  the  exposed  sur- 
faces and  hence  roasting  is  hastened  by  stirring  the  charge.  That  FeS 
which  is  near  the  surface  has  an  excess  of  air,  and  in  presence  of  silica 
which  acts  by  catalysis,  it  becomes  oxidized  thus: 

(6)  3FeS      +  HO   =     2SO2     +    Fe2O3  +  FeSO4, 
3X23,800  2X71,000       199,400       235,600= +505,600 

or  per  pound  of  sulphur,  5260  Cal.,  indicating  an  energetic  exothermic 
reaction.  Of  the  products  of  the  reaction  the  sulphur  dioxide  is  carried 
away  by  the  draft.  When  the  two  are  stirred  together,  the  Fe203  is  acted 
on  by  FeS  as  follows : 

(7)  FeS     +       10Fe203     =         7Fe3O4     +       SO2. 

23,800         10X199,400        7X265,800        71,000= +86,200 

Fe3O4  is  of  a  black  color;  when  the  ore  is  roasted  to  excess  the  resultant 
product  is  the  red  Fe2O3,  often  an  undesirable,  red-colored  product. 

As  the  charge  is  moved  toward  the  firebox,  the  iron  sulphate  produced 
as  in  (6)  begins  to  decompose  at  a  temperature  of  590°  C.  and  in  pres- 
ence of  cupric  oxide  reacts  as  follows: 

(8)  FeSO4    +    CuO    =     FeO    +       CuSO4 

235,600        37,200        66,400  182,600= -23800 

k 

That  is,  the  S03  given  off  by  FeS04  while  in  nascent  condition  is  taken 
by  the  CuO  to  form  its  sulphate,  the  reaction  being  an  endothermic  one. 
Such  of  the  FeS04  as  is  not  decomposed  by  the  cupric  oxide  is  broken  up 
by  the  heat  alone  as  follows : 

(9)  FeSO4    =     FeO    +     S03, 

235,600        66,400        9 1,800  =-77 ,400 

also  an  endothermic  reaction. 

At  a  slightly  greater  heat  (655°  C.)  the  cupric  sulphate,  formed  but  a 


CHEMISTRY  OF  ROASTING  91 

short  time  previously,  begins  to  decompose  and  at  a  dull  red  heat  the 
decomposition  of  cupro-cupric  sulphate  begins.  These  reactions  are  com- 
plete at  a  cherry  red  heat  (850°  C.)  up  to  this  temperature.  These  are  the 
reactions  of  a  sulphating  roast. 

At  850°  C.  the  zinc  and  lead  oxides,  reacting  on  the  copper  sulphate  now 
decomposing,  begin  to  be  changed  to  sulphates  thus: 

(10)  ZnO     +    CuSO4  =     ZnSO4  +     CuO, 
86,400         182,600        230,000  =   37,200  =-1800 

(11)  PbO     +  CuSO4    =    PbSO4  +     CuO. 
51,000         192,600        216,200        37,200=  -19,800 

These  last  four  reactions  are  endothermic  and  instead  of  aiding  the 
roasting  absorb  heat  as  the  result  of  the  reactions.  Fortunately  they 
take  place  in  the  hotter  part  of  the  furnace. 

As  the  charge  is  moved  nearer  the  fire  the  above  just-formed  sulphates 
decompose,  the  zinc  sulphate  more  readily  than  the  lead  sulphate,  and  SOs 
escaping. 

At  1050°  C.  (an  orange  heat)  copper  oxide  is  decomposed  into 
cuprous  oxide,  and  ferric  oxide,  losing  some  oxygen  becomes  FesO4. 

At  this  stage  the  ore  begins  to  fuse  if  it  contains  lead,  but  with  little 
lead  it  slightly  agglomerates,  with  much  it  fuses.  The  charge,  now  no 
longer  porous,  ceases  to  roast,  in  fact  it  is  hard  to  roast  such  an  ore  well. 
On  the  other  hand  a  zinc  ore,  free  from  lead,  can  and  should  be  brought  to 
a  high  finishing  heat  to  decompose  zinc  sulphate  and  to  eliminate  sulphur. 

To  decompose  completely  a  lead-bearing  zinciferous  ore  for  further 
treatment  in  a  blast-furnace,  the  following  procedure  is  successful.  After 
roasting  to  the  point  of  fusion,  the  ore  is  removed  to  a  reverberatory  melt- 
ing furnace.  Here  silicious  ore  is  added  and  the  whole  melted  at  a  high 
heat.  The  silica  reacts  on  the  lead  and  zinc  sulphates  thus: 

(12)  ZnSO4+Si02  =  ZnSiO3+SO3, 
(13) 


That  is,  the  sulphur  is  eliminated  as  a  sulphuric  anhydride,  leaving  a 
sulphur-free  silicate  of  both  metals.  In  the  blast-furnace  the  zinc  silicate 
enters  the  slag  as  such,  while  in  the  presence  of  fuel,  the  lead  is  reduced 
and  recovered. 

At  590°  C.  the  iron  sulphates,  formed  at  a  lower  temperature,  begin 
to  decompose. 

At  655°  C.  copper  sulphates,  formed  at  a  lower  heat,  begin  to  decom- 
pose. 

At  705°  C.  already-formed  cupro-cupric  sulphate  (CuSO4)  begins  to 
decompose. 


92  ROASTING 

At  850°  C.  copper  sulphates  are  entirely  decomposed,  and  when  steam 
is  present  the  maximum  amount  of  soluble  sulphate  (AgSCU)  is  formed. 

At  1050°  C.  copper  oxide  (CuO)  is  decomposed  to  Cu20. 

At  1100°  C.  ferric  oxide  (Fe2Os)  is  decomposed  to  the  next  lower  oxide 
Fe3O4. 

In  oxidizing  roasting  it  has  been  found  that  with  2  per  cent  862  by 
volume,  or  4.4  per  cent  by  weight  in  the  escaping  gases,  roasting  is  active. 
This  corresponds  to  46  Ib.  (570  cu.  ft.  at  sea  level)  per  pound  of  sulphur 
driven  off.  Calculating  this  for  a  16-ft.  MacDougall  roaster  treating  40 
tons  of  ore  in  twenty-four  hours,  and  roasting  it  from  35  per  cent  S  down 
to  7  per  cent  sulphur,  we  have  an  elimination  of  approximately  0.25  Ib. 
sulphur  per  second.  This  needs  142  cu.  ft.  of  free  air,  equal  to  284  cu.  ft.  of 
the  temperature  of  273°  C.,  that  of  maximum  chimney  discharge.  For  a 
velocity  of  20  ft.  per  second  in  the  stack  as  a  maximum  this  would  require 
an  area  of  14.2  sq.,  ft.  or  a  diameter  in  a  round  stack  of  4  ft.  3  in.  With 
an  excess  of  air  above  that  just  specified,  the  hearth  tends  to  cool  off; 
with  less,  roasting  proceeds  more  slowly,  so  that  at  4  .4  per  cent  SO2  in  the 
escaping  gases,  the  roasting  goes  actively,  at  8  per  cent  very  slowly  and 
at  9  per  cent  it  ceases  altogether. 

The  larger  the  charge  and  the  greater  its  thickness,  the  longer  is  the 
time  needed  to  complete  the  roast.  A  few  grams  are  roasted  in  a  half- 
hour  in  the  muffle,  and  in  twenty  hours  in  a  reverberatory  furnace,  while 
it  takes  weeks  to  roast  ore  in  the  pile. 

We  may  note  the  temperatures  of  the  reactions  that  occur  in  the 
reverberatory  furnace  as  follows : 

At  150°  C.  the  odor,  due  to  the  volatilization  of  some  of  the  loosely  held 
or  first  equivalent  of  sulphur,  can  be  detected. 

At  350°  C.  the  sulphur  of  the  sulphides  (particularly  of  pyrite)  begins 
to  burn  with  a  blue  flame. 

ROASTING  ORES  IN  LUMP  FORM— HEAP  ROASTING 

Heap  Roasting  has  the  advantage  that  it  can  be  used  at  the  first  instal- 
lation of  a  small  smelting  plant  in  a  new  district,  where  it  is  aimed  %)  avoid 
investment  in  an  expensive  roasting  plant.  It  requires  only  the  necessary 
site,  the  method  is  a  simple  one,  and  the  results  are  satisfactory.  On  the 
other  hand  in  a  large  plant,  where  from  10,000  to  50,000  tons  of  ore  or 
matte  are  in  process  of  treatment,  heap-roasting  may  cause  the  locking-up 
of  several  hundred  thousands  of  dollars  in  the  heaps. 

The  Chemistry  of  Heap  Roasting. — The  heat  generated  in  the  burning 
pile  volatilizes  sulphur  from  chalcopyrite,  and  where  there  is  insufficient 
air  some  of  it  escapes  from  the  top  of  the  pile  as  elemental  sulphur.  The 
remainder,  uniting  itself  according  to  the  equation:  S+ 02  =  862,  leaves 


HEAP  ROASTING  93 

as  sulphur  dioxide.  In  contact  with  heated  ore  it  is  further  changed  to 
sulphur  trioxide,  SO2+O  =  SO3. 

Air  has  access  to  the  exterior  of  the  lumps,  and  the  reaction  on  the  iron- 
copper  sulphides  is  as  follows : 

(14)  3FeS+10O  =  Fe3O4+3SO2. 

Site  of  Heap. — For  the  heaps  the  leaching  sites  should  be  nearly  flat, 
but  with  drainage  to  one  border,  layered  or  coated  with  clayey  slimes  and 
sprinkled  with  fuel-oil  for  tightness.  The  yard  for  a  large  site  should  have 
three  service  railroad  tracks,  two  for  the  green  or  unroasted  ore,  run 
along  outside  the  piles  170  ft.  apart,  the  roast-ore  tract  midway  between 
these.  This  leaves  room  for  two  rows  of  roast  heaps  each  60  ft.  wide, 
100  ft.  long  and  8  ft.  high,  to  hold  2500  tons. 

Heap  Building. — In  building  a  pile  or  heap  the  foundation  is  laid, 
usually  of  deadwood,  to  a  depth  of  12  to  18  in.,  the  surface  of  the  wood 
being  roughly  leveled.  At  intervals  of  10  ft.  are  left  flues  which  are 
filled  with  small  wood,  to  be  ignited  in  order  that  the  fire  may  penetrate 
rapidly  to  the  interior  of  the  heap  and  produce  a  more  uniform  com- 
bustion. Coarse  ore,  amounting  to  about  two-thirds  of  the  whole,  is  then 
piled  on  the  wood,  followed  by  a  layer  of  medium-sized  ore,  and  lastly  by 
fines  which  cover  the  top  and  sloping  sides  of  the  heap.  A  supply  of 
fines  is  kept  close  by,  so  that,  wherever  the  pile  is  burning  too  fast  in  any 
given  spot,  it  can  be  more  deeply  covered,  and  where  it  seems  dead,  the 
layer  can  be  opened  up  to  encourage  the  fire  to  that  spot. 

After  lighting,  the  wood  burns  out  in  about  sixty  hours,  leaving  the 
ore  in  vigorous  combustion.  The  pile  will  burn  for  three  or  four  months 
with  occasional  regulation  of  the  draft  as  above  described. 

During  the  roasting,  the  outer  portions  of  the  piles  become  reddish  in 
color,  due  to  the  oxidation  of  the  iron,  and  a  little 
sulphur  condenses  on  the  surface,  but  is  later 
driven  off.  The  raw  ore  may  be  specified  as 
averaging  23  per  cent  sulphur,  and  this  is  reduced 
to  10  or  12  per  cent.  Fig.  31  shows  a  cross-section 
of  a  lump  of  well-roasted  ore  containing  copper. 

The  copper  sulphide  is  not  so  easily  decom- 
posed, due  to  the  greater  affinity  of  copper  for    ^  70  _Roasted  Lump 
sulphur  as  compared  with  iron.     It  fuses  and  Ore 

accumulates  as  a  layer  beneath  the  oxidized  crust. 

As  this  crust  gets  thicker  so  also  does  the  layer,  until  finally  we  find  the 
lump  made  up  of  iron  oxide  with  a  center  of  copper  sulphides  of  a  bronze 
color,  and,  given  time,  even  this  sulphide  becomes  oxidized.  This  final 
condition  is  shown  in  Fig.  70. 

A  certain  amount  of  sintering  or  fusion  always  takes  place  and  parts  of 


94  ROASTING 

the  heaps  have  to  be  blasted  loose.  When  the  roasting  is  completed  the 
ore  is  dug  out  by  steam  shovel,  loaded  upon  cars  standing  on  the  center 
track  and  taken  to  the  blast-furnaces. 

Since  there  is  no  protection  from  the  weather  during  roasting,  and  since 
soluble  sulphates  of  copper  are  formed  during  the  operation,  a  portion  of 
the  valuable  metal  is  leached  out  by  rain  and  snow-water.  This  loss  is 
estimated  at  1  \  to  2  per  cent. 

The  cost  of  roasting  at  Ducktown,  Tenn.,  at  the  first  of  the  century  is 
said  to  be  42  cents  per  ton,  but  at  a  low  wage.  Peters  gives  a  cost  for  fuel, 
labor,  and  supplies  48.5  cents  per  ton  with  common  labor  computed  at 
$1.50  per  day.  Heap  roasting  may  be  done  by  contract  to  advantage. 
At  the  United  Verde,  Jerome,  Ariz.,  75  cents  per  ton  was  the  contract  price. 
The  Canadian  Copper  Co.,  in  1916,  roasted  its.  ore  on  a  large  scale  for  50 
cents  per  ton. 

Heap  Roasting  of  Matte. — Matte  can  be  well  roasted  in  lump  form,  but 
unlike  ore,  it  requires  two  or  more  burnings.  After  the  first  firing,  in  spite 
of  care,  matte  shows  but  little  the  change  it  has  undergone.  At  the  second 
burning,  using  a  larger  quantity  of  wood,  the  result  of  the  first  burning 
begins  to  show.  A  large  portion  of  the  twice-burned  material  is  found  to 
be  light  in  weight  and  porous,  and  to  contain  no  unburned  core.  In  fact 
the  thoroughness  of  the  roast  may  be  judged  by  feeling  of  the  lumps  with 
the  hand.  If  well-roasted  lumps  are  broken,  they  no  longer  show  the  raw 
core  at  the  center. 

The  bed  of  wood  can  be  prepared  for  matte  as  for  ore,  but  the  pile  is 
smaller,  being  only  12  ft.  square  by  6  ft.  deep,  with  a  single  chimney  at  the 
center.  The  broken  matte,  with  the  raw  fines  spread  over  it,  is  covered 
with  the  finer  portion  of  roasted  material.  The  burning  of  the  heap  lasts 
eleven  days,  and  when  ended,  it  is  taken  down,  and  the  imperfectly  roasted 
part  made  into  a  new  pile,  and  the  roasted  matte  sent  to  the  furnace.  It 
is  a  good  plan  in  constructing  the  new  pile  to  introduce  one  or  two  layers  of 
chips  or  bark,  for  a  reducing  effect  upon  impurities  like  arsenic,  and  for 
producing  a  more  uniform  heat  throughout  the  pile.  Finally,  after  this 
burning,  a  large  portion  suitable  for  use  can  be  sorted  out  and  the  part  still 
incompletely  burned  can  go  to  the  next  heap.  -4\ 

ROASTING  OF  ORES  IN  PULVERIZED  CONDITION 

This  work  is  done  in  single  or  multiple-hearth  furnaces.  The  ore  is 
spread  out  upon  the  hearth  or  floor  in  about  a  4-in.  layer,  exposed  to  the 
action  of  flame  and  air. 

If  not  already  fine  enough,  it  is  crushed  so  fine  that  the  particles  at 
the  end  of  the  time  given  for  roasting  show  no  unburned  core  or  center. 
An  ore,  mainly  iron  pyrite,  decrepitates  in  roasting,  hence  is  fine  enough 


ROASTING  IN  FURNACES  95 

if  of  two-  or  three-mesh  size.  Many  ores  and  matte  need  crushing  to  four- 
or  six-mesh  size.  An  ore  of  blende  or  galena  is  compact  and  when  to  be 
roasted  (especially  when  "  dead  roasted  "  so  that  no  sulphur  is  left)  had 
better  be  ground  to  10-mesh  size.  Where  the  ore  is  finely  ground  for  subse- 
quent leaching  this  may  be  done  before  roasting.  It  is  a  good  plan,  how- 
ever, to  grind  to  a  coarse  size  for  roasting,  and  to  regrind  the  roasted  prod- 
uct as  fine  as  desired  for  after  treatment. 

In  the  various  furnaces  advantage  is  taken  of  the  heat  developed  by  the 
burning  of  the  sulphides,  especially  in  the  compact  multiple-hearth  roasters. 
If  the  percentage  of  sulphur  is  high,  this  is  often  enough  to  supply  the 
required  heat  (after  combustion  has  once  been  started)  without  the  aid  of 
extraneous  fuel.  Thus  in  the  MacDougall  roaster,  after  the  furnace  and 
ore  has  been  sufficiently  heated,  a  content  of  25  to  30  per  cent  sulphur 
ensures  the  continuance  of  roasting. 

The  various  mechanical  roasters  treat  ore  cheaply,  but  for  ores  con- 
taining much  lead,  which  agglomerate  or  sinter,  they  do  not  work  well. 
With  a  slight  accession  of  heat  above  the  normal,  caused  by  lack  of  care 
in  firing,  the  ore  is  liable  to  agglomerate,  and  eventually  to  stick  to  the 
hearth,  stopping  the  movement  of  the  rabbles.  When  this  becomes 
serious  a  stout  flat  bar  of  iron,  attached  to  one  of  the  rabble  arms  in  place 
of  a  rabble  blade,  may  plow  up  these  accretions,  and  by  setting  it  in  dif- 
ferent positions  on  the  arm  the  hearth  may  be  finally  cleared.  The  device 
has  not  proved  entirely  successful.  In  the  hand  reverberatory  roaster  the 
hearth  is  accessible,  and  when  the  accumulation  builds  upon  the  hearth 
it  may  be  removed  by  aid  of  cutter-bars  and  a  hammer.  If,  however,  it 
is  sufficient  to  rough-roast  such  an  ore,  reducing  the  sulphur  content,  no 
more  than  from  10  to  13  per  cent,  then  such  agglomeration  need  not  be 
feared.  Hand  roasters  work  well  upon  ores  that  need  a  high  finishing  heat 
suited  to  breaking  up  or  decomposition  of  the  sulphates,  as  in  the  roasting 
of  zmc  ores  or  of  galena.  The  objection  to  such  roasting  is  that  it  is  costly. 

THE  LONG-HEARTH  REVERBERATORY  ROASTER 

In  this  furnace  the  charge  is  put  in  and  removed  at  intervals. 
These  furnaces  may  be  distinguished  from  the  reverberatory  furnace 
used  for  melting  by  the  relatively  small  grate  area  and  by  the  fact 
that  the  hearth  is  flat  and  at  the  level  of  the  door  sills.  The  hearth 
may  be  10  ft.  wide  by  36  ft.  long,  divided  as  shown  on  the  plan,  Fig.  71, 
into  three  hearths  with  a  drop  of  2  in.  between.  (Large  furnaces  are 
built  70  ft.  long  with  five  hearths  14  ft.  wide.)  The  length  of  a  hearth 
for  a  reverberatory  roaster  should  accord  with  the  percentage  of  sulphur 
that  the  ore  contains,  and  in  consequence  the  heat  developed  by  it  in  roast- 
ing. Without  the  aid  of  the  heat  developed  as  the  result  of  the  burning  of 


96 


ROASTING 


I 


•a 

I 


SECTION  O-D 

FIG.  71. — Reverberatory  Roasting-furnace  (sections). 


MECHANICAL  ROASTING  FURNACES  97 

the  sulphur,  the  fire  would  not  maintain  sufficient  heat  to  roast  ore  25  ft. 
from  the  fire-bridge.  An  ore  containing  10  per  cent  sulphur  can  be  roasted 
to  good  advantage  in  a  furnace  having  a  single  hearth  of  15  ft.;  when 
15  per  cent  sulphur  is  present  we  may  add  another  hearth,  bringing  the 
length  to,  say,  30  ft.;  a  20  per  cent  ore  would  work  rapidly  in  a  three- 
hearth  furnace;  an  ore  of  29  to  33  per  cent  would  do  well  on  a  four-  or 
five-hearth  furnace. 

To  furnish  draft  for  a  stack  or  chimney  (see  the  plan  Fig.  71),  28  in. 
diameter  inside  by  65  ft.  high  will  be  sufficient. 

Operation. — Into  the  thoroughly  hot  furnace  (the  slide  of  the  charge- 
hopper  being  withdrawn)  a  charge  of  2000  Ib.  pours  in  a  conical  heap  on  the 
first  hearth  and  is  there  spread  by  a  man  on  each  side  using  a  paddle. 
(This  has  an  iron  pipe-handle  12  ft.  long  with  a  blade  6  in.  wide  by  18  in. 
long.)  Here  the  ore  remains  for  four  hours  being  stirred  every  half-hour 
with  a  rabble,  a  hoe  having  a  blade  of  6  by  10  in. 

By  means  of  the  paddles  it  is  then  moved  down  and  spread  out  on 
the  second  hearth,  while  the  first  hearth  receives  a  fresh  charge  from  the 
hopper. 

Again  at  the  end  of  four  hours  both  charges  are  moved  toward  the  fire, 
and  another  charge  dropped  on  hearth  No.  1.  Thus  all  the  hearths 
become  covered  with  ore.  At  the  expiration  of  the  four-hour  period  the 
first  charge  now  on  the  last  hearth  is  withdrawn  through  a  square  discharge 
hole  seen  near  the  fire-bridge  in  the  figure.  It  drops  into  a  wheelbarrow 
set  beneath.  This  ore  has  thus  been  under  the  action  of  the  fire  for 
twelve  hours,  and  for  the  charge  specified  we  compute  an  output  of  6  tons 
daily  of  raw  ore  or  as  much  as  5  tons  of  roasted  ore  or  calcine. 

MECHANICALLY  OPERATED  ROASTING  FURNACES 

These  may  be  classified  as  follows: 

(a)  The  revolving  cylinder  furnace  with  the  axis  horizontal  or  inclined 
toward  the  discharge  end.  An  example  of  such  a  furnace  is  the  Bruckner 
cylinder  roaster,  having  a  horizontal  axis,  the  cylinder  7  ft.  diameter  by 
25  ft.  or  more  long.  The  ore  is  roasted  in  batches  or  charges.  It  is 
charged  and  discharged  through  manholes.  A  charge  of  40  tons  may  take 
three  days  to  roast.  The  Oxland,  the  White-Howell,  and  the  Argall  are 
examples  of  cylinder  furnaces  with  the  axis  inclined,  causing  the  ore  grad- 
ually to  travel  from  the  feed  end  to  the  lower  or  discharge  end,  as  the  fur- 
nace revolves.  The  Oxland  and  the  White-Howell  are  single-cylinder 
furnaces.  In  the  Argall  four  are  united  in  one. 

(6)  The  mechanically  rabbled  reverberatory  furnaces  having  a  con- 
tinuous feed  and  discharge.  Of  these  the  Brown-O'Harra,  the  Ropp,  the 
Edwards,  the  Merton,  the  Wethey,  and  the  Hegeler  have  straight  hori- 


98 


ROASTING 


zontal  hearths  and  so  are  called  straight-line  furnaces.  The  first  four 
have  single-hearths,  the  fifth  has  two  superimposed  hearths  and  the 
Hegeler  is  a  multiple-hearth  furnace. 

Another  variation  of  horizontal  hearth  furnace  has  the  hearth  curved 


as  in  the  Brown  horseshoe  furnace,  or  circular,  as  in  the  Pierce  turret  fur- 
nace. The  ore  in  either  case  having  made  the  circuit  of  the  hearth  is 
discharged.  The  rakes  or  rabbles  are  but  part  of  the  time  in  the  furnace 
in  order  that  they  may  have  time  to  cool.  The  Brown  horseshoe  has  but 
a  single  hearth  while  the  Pierce  turret  may  be  single  or  double. 


THE  WHITE-HOWELL  FURNACE 


99 


A  type  of  furnace  once  used  in  the  chloridizing  roasting  of  silver  ores 
was  the  Stetefeldt  where  the  ground  ore  was  showered  down  a  shaft. 

The  White-Howell  Cylinder  or  Furnace. — Fig.  72  is  a  longitudinal  ele- 
vation of  this  furnace.  It  consists  of  a  cylinder,  50  in.  inside  diameter  by 
34  in.  long,  set  at  an  inclination  of  2\  per  cent  supported  on  friction-rollers 
carried  on  the  driving  shaft.  At  one  end  is  the  firebox,  at  the  other  a 
dust-chamber  which  connects  by  a  flue  to  the  stack.  The  hotter  end  of 
the  cylinder,  near  the  firebox,  is  of  larger  diameter,  to  permit  of  its  being 
lined  with  brick,  thus  leaving  the  cylinder  of  uniform  interior  diameter 
throughout.  Projecting,  longitudinal,  firebrick  ledges,  set  spirally,  raise 
the  ore  and  shower  it  back  through  the  flame  as  the  cylinder  revolves,  so 
as  to  roast  it  more  rapidly.  The  unlined  part  for  the  same  reason  is  fur- 
nished with  longitudinal,  cast-iron,  projecting  shelves.  Ore  is  fed  at  the 
flue-end,  by  means  of  a  screw-feed  (see  Fig.  267),  and  when  dropped  into 


FIG.  73.— Cylinder  Dryer. 

the  revolving  cylinder,  travels  along,  discharging  at  the  firebox  end.  Just 
before  it  reaches  the  firebox  it  passes  out  from  the  cylinder  to  a  brick 
chamber  below,  and  is  withdrawn  from  that  when  cool.  The  furnace 
makes  much  flue-dust.  It  is  used  chiefly  for  chloridizing  roasting,  upon 
ores  containing  but  little  sulphur,  and  has  a  capacity  of  50  tons  per  twenty- 
four  hours  for  low-sulphur  ores. 

For  an  ore-drier  a  furnace  quite  like  this  is  employed,  as  shown  hi 
Fig.  73. 

The  Edwards  Roasting  Furnace. — This  is  a  single-hearth  reverberatory 
furnace  with  hearth  dimensions  57  ft.  long  by  6  ft.  wide.  Fig.  74,  in  plan, 
shows  a  portion  at  the  firebox  end,  the  feeding  mechanism  and  the  cooling 
floor  in  section.  The  elevation  shows  the  side,  constructed  like  a  plate- 
iron  beam,  the  stirring  mechanism  and  the  conveyor  for  transferring  the 
roasted  ore  to  the  cooling-pit.  Fig.  75  is  a  transverse  section  of  the  hearth 
showing  the  details  of  the  stirring  mechanism.  The  slope  of  the  furnace 
can  be  changed  a  little  by  tilting.  This  regulates  the  rate  of  travel  of  the 


100 


ROASTING 


THE  EDWARDS  ROASTING,  EURJSL\C£ \ 


101 


ore  through  the  furnace;  but  for  a  given  kind  of  ore,  this  slope,  once 
determined,  is  not  again  changed.  The  furnace  has  a  slope  of  \  in.  per 
foot  toward  the  discharge  or  firebox  end.  The  stirring  and  propulsion  of 
the  charge  are  effected  by  means  of  rabbles  fixed  to  vertical  shafts,  as  shown 
in  the  elevation  of  Fig.  75,  and  in  the  plan  of  Fig.  74.  These  rabble-shafts 
.make  one  revolution  in  sixty  to  ninety 
seconds.  The  rabbles  at  the  firebox 
end  are  water-cooled,  and  this  is 
found  especially  necessary  where  a 
high  finishing  heat  is  needed.  The 
blades  or  plows  of  the  rabbles  can  be 
easily  replaced  through  the  doors 
adjacent  to  them.  The  figure  indi- 
cates the  hearth  as  broken  away,  at 
the  discharge  end,  to  show  two  of  the 
rabbles  in  plan.  The  last  rabble 
sweeps  the  roasted  ore  into  the  dis- 
charge shoot,  and  the  push-conveyor 
then  moves  the  ore  to  the  cooling- 
pit.  The  bottom  of  the  conveyor 
trough  is  furnished  with  slides,  by 


FIG.  75. — Edwards  Roasting  Furnace 
(Tran verse  Section). 


means  of  which  the  ore  can  be  dropped  at  any  desired  point  on  the 
cooling-floor.  The  ore  is  fed  to  the  furnace  from  the  feed-hopper,  by  an 
endless-screw  conveyor  which  discharges  into  a  feed-opening  in  the  roof 
of  the  furnace.  The  smoke  is  carried  off  by  a  flue.  The  furnace  takes 
1  H.P.  to  operate,  and  has  a  daily  capacity  of  25  tons  on  sulphide  ore  of 
30  to  35  per  cent  sulphur.  The  roasted  ore  contains  3  to  8  per  cent  of 
sulphur.  The  moving  parts  are  durable,  and  the  furnace  has  proved 
efficient  in  practice.  Large  installations,  of  the  duplex  type  with  a  double 
instead  of  a  single  row  of  rabbles,  and  of  hearth-dimensions  120  by  12  ft., 
have  been  built  for  a  daily  capacity  of  60  tons.  These  furnaces  do  not 
have  the  tilting  hearth. 

Besides  these  we  find  circular  revolving-hearth  roasters,  as  the  Brunton 
and  the  Spirlet.  The  Brunton  furnace  is  used  in  the  roasting  of  arsenic- 
bearing  flue-dust.  The  Spirlet  is  used  hi  blende  roasting;  the  hearth 
revolves  and  the  rakes  or  blades  are  fixed  in  the  roof  above. 

The  Multiple-hearth  Furnace  Type. — A  group  of  mechanical  roasters 
of  the  circular,  multiple-hearth  type  are  to-day  most  used.  These  are  the 
MacDougall,  the  Wedge,  and  the  Herreshoff .  Of  them  we  will  fully  describe 
the  MacDougall  and  the  Wedge  roasters. 

The  MacDougall  Roasting  Furnace. — There  are  several  kinds  of  furnaces 
of  this  type.  Among  these  are  the  Herreshoff  and  the  Wedge.  The 
MacDougall  furnace  as  manufactured  by  the  Allis-Chalmers  Co.  is  shown 


102 


>  ROASTING 


in  sectional  elevation  in  Fig.  77.  It  is  a  vertical,  cylindrical  furnace,  18  ft. 
diameter,  with  six  arched  hearths,  over  which  travel  rabbles  which  stir 
and  move  the  ore  gradually  toward  the  drop-openings  through  the  floor 
of  each  hearth,  situated  alternately  at  the  center  and  at  the  periphery.  A 
central  shaft  is  provided,  carrying  six  radial  rabble-arms  (three  of  these 

are  hidden  by  the  shaft  in  the 
illustration),  provided  with 
rabble-blades  set  at  an  angle 
on  the  arm. 

The  rabble-blades  on  the 
even-numbered  hearths  are  so 
set  as  to  push  the  ore  in  a 
spiral  path  toward  the  pe- 
riphery; the  odd-numbered 
ones  toward  the  center.  The 
ore,  fed  continuously  into  the 
furnace  from  a  cylindrical 
hopper  shown  above  and  at 
the  right,  Fig.  76,  drops  upon 
the  upper  hearth  near  its 
outer  edge.  The  rabble-blades 
of  that  hearth  stir  and  move 
the  ore  gradually  toward  the 
central  drop-opening  where  it 
falls  to  hearth  No.  2.  The 
rabbles  of  this  hearth  again 
stir  and  move  it  to  the  outer 
drop-openings,  through  which 
it  falls  to  hearth  No.  3.  The 
ore  advances  by  this  means 
until  it  reaches  the  lower  hearth,  where  an  opening  at  the  periphery  gives 
it  exit  to  a  receiving-hopper,  shown  beneath  the  hearth,  from  which  it  is 
drawn  into  a  car  as  required. 

A  high-sulphide  ore  roasts  by  its  own  heat  when  the  furnace  is  iii  full 
operation.  The  ore  fills  the  hearth  to  the  level  of  the  blades,  and  is  spread 
out  evenly  by  them.  On  the  upper  hearth,  as  the  ore  moves  toward  the 
central  opening,  it  becomes  dry  and  hot,  and  when  dropped  upon  hearth 
No.  2,  begins  roasting.  On  hearth  No.  3,  the  ore  roasts  freely,  emitting 
sparks  and  forming  sulphates.  On  hearth  No.  4  no  sparks  are  seen,  and 
the  ore  has  attained  its  highest  temperature.  On  hearth  No.  5  the  ore 
looks  less  bright;  and  on  No.  6,  especially  at  the  discharge,  it  has  become 
cooler. 

The  air  for  oxidation  is  admitted  by  side  doors,  mostly  those  of  the  lower 


FIG.  76.— Six-hearth  MacDougall  Roasting 
Furnace. 


THE  MAcDOUGALL  ROASTING  FURNACE  103 

hearths.  The  gas  and  dust,  passing  up  through  the  drop-openings,  are 
drawn  through  the  horizontal  main  flue.  In  starting,  the  furnace  is  heated 
to  the  kindling  temperature  of  the  ore  which,  if  rich  in  sulphur,  burns  by  its 
own  heat,  without  the  aid  of  fuel.  If  the  sulphur  content  is  low,  addi- 
tional heat  is  supplied  by  one  or  more  external  fireplaces,  near  the  bottom 
of  the  furnace. 

To  protect  the  rabble-arms  from  the  intense  heat  they,  and  likewise  the 
central  shaft,  are  water-cooled.  The  cooling-water  is  forced  down  the 
9-in.  hollow,  central  shaft  in  a  3-in.  pipe  to  a  point  near  the  bottom,  and 
out  to  the  ends  of  the  arms  in  1-in.  pipes.  It  then  returns  up  the  annular 
space  between  the  3-in.  pipe  and  the  hollow  shaft,  and  discharges  at  the 
top  through  two  spouts  into  a  launder.  The  furnace  is  18J  ft.  high  by  18  ft. 
diameter,  and  has  a  total  hearth-area  of  1600  sq.  ft.  The  structure  is  sup- 
ported on  columns  to  give  room  below  for  the  hopper  and  the  car  into 
which  the  roasted  ore  is  discharged.  The  shell,  made  of  f-in.  plate-steel 
is  lined  with  9  in.  of  brick-work.  The  rabble-arms  consume  1J  to  2  H.P. 
and  make  one  revolution  in  1 J  minutes. 

A  furnace  treats,  in  twenty-four  hours,  65  tons  of  sulphide  ore  of  35 
per  cent  sulphur,  reducing  it  to  7  per  cent.  About  4  per  cent  flue-dust  is 
made ;  and  the  ore  itself  contains  more  ferric  oxide,  and  is  lighter  and  more 
porous  than  if  treated  in  a  hand-reverberatory  roaster.  The  cost  of 
roasting  such  ore  is  approximately  35  cents  per  ton,  which  is  the  lowest 
figure  thus  far  known  for  any  furnace.  The  compact  form  of  the  furnace 
reduces  radiation  to  a  minimum  and  permits  roasting  with  little  or  no  fuel. 
Taking  capacity  into  consideration,  the  furnace  is  one  of  moderate  price, 
and  one  that  costs  little  to  keep  in  repair. 

Of  the  two  revolving-hearth  furnaces,  the  Holthoff  and  the  Raymond, 
the  latter  has  some  popularity  for  the  preliminary  roasting  of  ores  for 
blast  or  pot-roasting,  the  powdered  ore  being  showered  down  a  vertical 
shaft  or  tower  and  coming  in  contact  with  an  upward  flame  from  a  fire- 
box. An  objection  to  its  use  is  that  much  flue-dust  is  made. 

Fig.  76  is  a  perspective  view  of  the  enclosed  type  of  MacDougall  roasting 
furnace,  where,  in  order  to  permit  a  firebox  below  the  lowest  hearth  the 
driving  mechanism  has  been  transferred  above  the  furnace.  The  firebox 
enables  the  hearths,  and  especially  the  lower  one,  to  be  heated.  By  its 
use  the  ore  may  be  roasted  to  a  lower  percentage  in  sulphur. 

In  Fig.  77  we  give  a  plan  and  sectional  elevation  of  a  MacDougall 
roaster.  At  the  center  drop  holes,  a  plate  on  the  rabble  arms  hold  up  the 
ore  close  to  the  opening  so  that  the  gases  do  not  pass  upward  there.  The 
outer  drop-holes  are  similarly  sealed.  Thus  the  gases  must  pass  upward 
by  the  gas  passageways,  and  are  thus  free  from  flue  dust. 

The  Wedge  Roasting  Furnace. — Fig.  78  is  a  sectional  elevation  of  a 
seven-hearth  furnace  as  constructed  for  oxidizing  roasting.  If  the  ore 


104 


ROASTING 


FIG.  77. — MacDougal  Roasting  Furnaces  (sections). 


THE  WEDGE  ROASTING  FURNACE 


105 


contains  moisture,  this  is  dried  out  upon  the  top,  called  the  drier-hearth. 
The  central  hollow  shaft  A  is  covered  outside  with  tiles  to  protect  it  from 
the  furnace  heat.  It  carries  at  each  hearth  two  opposite  rabble-arms  which 
are  water-cooled  by  small  feed  and  discharge  pipes  leading  down  from 
above.  Individual  pipes  from  the  water  pan  E  pass  down  the  shaft  to  the 


FIG.  78. — Sectional  Elevation  of  Wedge  Roaster. 

rabble  arms.     The  hearths  are  enclosed  hi  a  steel  shell  22J  ft.  diameter. 
The  outlet  for  the  sulphur-bearing  gases  is  at  F. 

The  driving  pulley  P  through  a  train  of  spur  gears  communicates  motion 
to  a  bevel  pinion  on  the  end  of  a  horizontal  shaft,  and  this  meshes  into  the 
master  gear  G,  at  the  foot  of  the  hollow  shaft  A.  This  same  gear  has  a 
turned  raceway  at  its  edge  supported  by  rollers  that  carry  the  whole  weight 
of  the  shaft  and  the  sixteen  rabble  arms. 


106 


ROASTING 


THE  WEDGE  MECHANICAL  FURNACE 


WEDGE  MECHANICAL  FURNACE  CO, 

•  P  H  1 1_  A  O  E  L.  P  M  I  A.  • 


FIG.  79. — Elevation  of  Wedge  Roaster. 


Automatic 
Tripper 


FIG.  80.— Section  of  Roaster  Plant. 


THE  BRUNTON  ROASTING  FURNACE 


107 


The  ore,  fed  at  the  outer  edge  of  the  top  or  drier  hearth,  works  down 
under  the  central  plate  upon  hearth  No.  1.     It  passes  across  the  hearth  to 


YS///S/////////////////////////     ///,  ////////////' 

SECTION  C-C  SECTION  D-D 


HALF  SECTION  B-B 


FIG.  81 . — Brunton  Roasting  Furnace. 

the  outer  drop-holes  of  hearth  No.  2.     On  hearth  No.  3  the  ore  has  drop- 
holes  other  than  the  central  opening  where  the  gases  rise  so  there  is  less 


108  ROASTING 

flue  dust  made  and  so  on.  The  lower  hearth  has  a  peripheral  drop-hole 
for  the  discharge  of  the  calcine.  Fig.  78. 

The  method  of  installation  of  a  roaster  building  containing  multiple- 
hearth  roasters  is  shown  in  Fig.  80.  Crushed  ore  from  the  sampling  and 
grinding  building  is  carried  to  its  top  story  by  an  incline  belt  conveyor  and 
is  delivered  to  the  roaster  feed-hopper  by  means  of  an  automatic  tripper  on 
the  horizontal  20-in.  conveyor  belt. 

The  Brunton  Roasting  Furnace. — In  Fig.  81  we  show  a  plan  and  eleva- 
tion of  this  furnace,  also  sections  of  the  firebox.  This  has  a  revolving  hearth 
and  three  sets  of  rabbles,  being  iron  blades  passing  through  the  roof. 
These,  set  at  an  angle,  stir  and  move  the  ore,  which  is  fed  through  the  roof 
at  the  center  to  the  discharge  chute  at  the  right.  The  hearth  is  carried  on  a 
cast-iron  frame  fixed  to  a  vertical  shaft  and  is  driven  by  a  large  worm- 
wheel  and  a  worm  on  the  horizontal  driving  shaft.  There  are  two  coal- 
fireboxes  kk,  the  flames  entering  at.  the  side  of  the  hearth  and  the  mingled 
fire  gases  and  arsenical  fumes  escaping  at  flue  m. 

ROASTING  OF  MATTE 

The  term  "  roasting  "  is  applied  also  to  a  method  of  treating  copper 
matte  in  a  reverberatory  furnace  in  large  pieces,  upon  which  an  oxidizing 
flame  is  allowed  to  play.  Such  masses  slowly  melt  and  are  acted  on  by  the 
air,  whereby  a  part  of  the  material  becomes  oxidized  or  roasted  sufficiently 
for  the  next  operation.  As  compared  with  ordinary  roasting  this  is  slow, 
and  the  method  is  one  but  little  used. 

Copper-bearing  matte  to  be  subjected  to  an  ordinary  oxidizing  roast 
must  be  crushed  at  least  to  4-mesh  zize.  Matte  from  the  silver-lead 
smelting  to  be  roasted  in  a  reverberatory  furnace  of  the  kind  shown  in 
Fig.  63,  needs  a  different  treatment  from  that  given  to  ore.  This  kind 
of  matte  contains  but  20  per  cent  sulphur,  and  does  not  take  fire  like 
pyrite  ore,  but  must  have  a  high  finishing  heat  to  expel  the  sulphur.  Such 
matte  is  considered  well  roasted  when  it  contains  4  per  cent  sulphur. 
Ores  low  in  lead  can  easily  be  roasted  to  2  to  3  per  cent  sulphur,  while 
galena,  when  roasted,  still  contains  5  to  6  per  cent  when  drawn  fro|n  the 
furnace.  Like  matte,  galena  starts  burning  slowly,  and  must  be  roasted 
slowly,  for  rapid  heating  causes  it  to  sinter  and  thus  stops  further  roasting. 
Typical  leady  matte  contains  metals  and  sulphur  as  shown  in  the  sub- 
joined table. 

The  roasted  low-grade  matte  contains  23  per  cent  oxygen.  This 
explains  why  it  does  not  lose  weight  in  roasting.  Pyrite  ores  of  20  to 
30  per  cent  sulphur,  on  the  contrary,  easily  lose  15  per  cent  hi  weight. 

Losses  in  Roasting. — Such  loss  depends  upon  the  extreme  to  which  the 
roasting  is  carried  as  well  as  upon  the  nature  of  the  ore.  When  ore  is  so 


ROASTING  COSTS 


10$ 


Raw,  Low 
Grade. 
Per  Cent. 

Roasted, 
Low  Grade. 
Per  Cent. 

Raw, 
Shipping. 
Per  Cent. 

Cu               

4.62 

4   12 

42  30 

Fe                            

53    11 

52  41 

20  00 

s                     

26  87 

6  13 

17  89 

Pb                                         

10  66 

10  49 

9  06 

95.26 

73.15 

89.25 

roasted  that  it  is  not  sintered  at  the  final  high  temperature,  the  lead  lost 
averages  2.5  per  cent,  but  no  loss  of  silver  occurs.  When  the  tempera- 
ture is  carried  higher,  and  the  ore  is  agglomerated,  the  loss  is  slightly 
higher.  When  fused  it  may  reach  15  to  20  per  cent  of  the  lead  and  2  to  5 
per  cent  of  the  silver.  Of  the  gold  little  is  lost  in  oxidizing  roasting. 

CAPACITY  OF  FURNACES  AND  COST  OF  ROASTING 

These  depend  upon  the  surface  exposed  to  the  oxidizing  influences  and 
upon  the  quantity  of  sulphur  contained  in  the  ore  in  hand  reverberatory 
roasting.  Silicious  ore,  containing  \  to  3J  per  cent  sulphur,  requires  13  to  15 
sq.  ft.  of  hearth-area  per  ton  of  ore  roasted  per  twenty-four  hours.  Matte 
containing  20  to  25  per  cent  sulphur,  when  it  is  necessary  to  reduce  the 
sulphur  content  to  4  per  cent,  needs  45  sq.  ft.  hearth-area;  copper  sulphide 
ore,  roasted  to  7  per  cent  in  preparation  for  smelting,  requires  33  to  35 
sq.  ft.  For  roasting  iron-sulphide  concentrate,  which  carries  35  to  45 
per  cent  sulphur,  down  to  3  to  10  per  cent  sulphur,  55  to  60  sq.  ft.  hearth- 
area  is  needed. 

Roasting  Costs. — In  1910  to  1913  ore-roasting  in  heaps,  at  Jerome,  Ari- 
zona, cost  80  cents  per  ton,  including  general  expense.  Ore-roasting  in  stalls 
cost  50  cents  per  ton.  For  reverberatory  roasting,  in  long,  hand-rabbled 
furnaces  the  lowest  price  attainable  on  copper  ores  was  $1.50,  with  an  aver- 
age of  $1.81  per  ton.  For  roasting  lead-bearing  ores,  $1.75  is  a  moderate 
cost,  and  from  this  the  cost,  when  all  items  are  included,  may  rise  to  $2.25 
per  ton.  The  Allen-O'Harra  automatic  furnace,  having  two  straight  hearths 
each  94  by  9  ft.,  and  resembling  the  Wethey  furnace,  roasts  45  to  50  tons 
daily  at  a  cost  of  78  cents  per  ton.  The  Wethey  furnace,  of  the  type  having 
four  hearths,  each  65  by  10  ft.,  the  roasting  proceeding  on  all  the  hearths, 
roasts  90  tons  daily  to  5  to  6  per  cent  sulphur,  at  a  cost  of  98  cents  per  ton. 
The  16-ft.  MacDougall  furnace  (Herreshoff  type),  having  five  hearths, 
14|  ft.  diameter,  and  a  total  area  of  830  sq.  ft.,  roasts  33  to  35  tons  daily  to 
7  per  cent  sulphur  at  a  cost  of  50  cents  per  ton.  The  Bruckner  roasting 
cylinder,  8J  ft.  diameter  by  22  ft.  long,  takes  a  charge  of  20  tons  (10  tons 
daily) ,  and  in  forty-eight  hours  roasts  it  to  4  per  cent  sulphur  at  a  cost  of 
80  cents  per  ton. 


110  ROASTING 

It  will  be  noticed  that  the  low  cost  of  roasting  in  some  of  these  furnaces 
is  due  to  their  needing  no  fuel  after  coming  into  full  operation.  To  obtain 
this  effect  such  furnaces  have  several  hearths,  and  are  compact.  On 
account  of  this  compactness  they  lose  but  little  heat  by  radiation. 

The  above  roasting  costs  were  for  the  period  1910  to  1913,  but  these 
figures  must  be  doubled  for  present  conditions. 

BLAST-  OR  POT-ROASTING  OF  ORES 

Both  lead  and  copper  ores  are  treated  by  blast-  or  pot-roasting,  though 
the  method  was  at  first  intended  for  lead-bearing  ores,  especially  for  galena. 
I  have  already  mentioned  the  difficulty  of  roasting  galena  by  the  old 
method,  in  the  reverberatory  furnace;  but  by  pot-roasting,  it  can  be  so 
treated  as  to  remove  most  of  its  sulphur,  with  less  loss  by  volatilization. 

Treatment  of  Galena. — By  the  Huntington-Heberlein  process,  called 
also  the  "  H  and  H  process,"  the  galena-bearing  ore  is  given  an  incomplete, 
rather  rapid  roast,  to  reduce  the  amount  of  sulphur  to  12  to  14  per  cent. 
The  product  from  the  roaster  is  mixed  with  a  certain  proportion  of  lime- 
stone and  silicious  ore,  wet  down,  and  charged  into  a  hemispherical  cast- 
iron  pot  8J  ft.  diameter  by  4  ft.  deep,  having  a  capacity  of  8  to  10  tons  as 
shown  in  Fig.  82.  Within  the  pot,  and  forming  a  false-bottom,  is  placed 
a  circular  arched  plate  perforated  with  f-in.  holes  to  admit  air  to  the  charge 
under  pressure.  Upon  the  false-bottom  is  scattered  a  wheelbarrow-load 
of  ashes,  then  a  carload  (one  ton)  of  hot  ore  from  the  roaster.  On  this  is 
dumped  8  tons  of  charge  wet  to  about  6  per  cent  moisture.  Air,  under  the 
pressure  of  a  few  ounces,  is  admitted  beneath  the  false-bottom,  and  coming 
up  through  the  hot  ore,  it  produces  a  burning-temperature  and  starts  the 
combustion  of  the  charge.  The  heat  gradually  ascending  to  the  top,  the 
charge  becomes  red-hot,  and  S02  and  SOs  escape.  At  the  end  of  the 
roasting,  which  lasts  sometimes  sixteen  hours,  there  remains  only  3  to  5  per 
cent  sulphur  if  the  charge  is  properly  burned.  The  pot  is  now  inverted  to 
discharge  the  contents,  and  this  falls  out  in  an  agglomerated,  red-hot 
mass.  It  is  broken  to  a  size  suited  to  subsequent  treatment  in  the  blast- 
furnace. ^ 

The  Dwight-Lloyd  Machine. — Blast-roasting  in  pots  has  several  dis- 
advantages: the  ore  is  exposed  for  a  long  time  to  the  hot  gases  and  this 
leads  to  a  loss  of  metal;  the  process  is  intermittent;  the  charge  needs  con- 
stant attention  either  in  charging,  discharging  or  blowing;  the  amount 
of  fine  is  apt  to  be  considerable  and  this  must  be  re-treated ;  the  ore  is  not 
evenly  sintered;  finally,  it  is  expensive  to  break  up  the  sintered  mass, 
whether  by  hand  or  by  power.  These  disadvantages  appear  to  be  over- 
come by  the  Dwight-Lloyd  sintering  process,  especially  by  the  use  of  their 
endless-chain  machine,  28  ft.  total  length,  illustrated  in  Fig.  83.  As 


THE  DWIGHT-LLOYD  BLAST-ROASTER 


111 


compared  with  j»ot-roasting,  and  particularly  with  roasting  furnaces,  this 
machine  occupies  but  little  room. 

The  endless-chain  carries  a  train  of  pallets.     Each  pallet  is  in  fact  a 
perforated  grate  having  two  edges  upturned  4J  in.     The  joints  between  the 


Pedestal:  6" round  at  liase, 
4" round  at  top,  with  2"hole 

FIG.  82. — Details  of  Construction  of  Blast-roasting  Pot. 

pallets  and  between  the  pallets  and  the  suction-box  are  close  and  fit  snugly 
by  planed  edges.  At  the  ends  of  the  suction-box  a  planed  dead-plate,  over 
which  the  pallets  glide,  serves  to  make  the  joint  tight  there.  After  the 
pallets  leave  the  suction-box  their  four  wheels  transfer  their  weight  to  the 
rails.  At  the  delivery-end,  where  a  car  is  stationed,  the  pallets  are  slightly 


112 


ROASTING 


raised,  fracturing  the  sintered  coke,  then  drop  one  by  one,  striking  the 
pallet  next  below  on  the  circuit,  while  the  sintered  ore  is  jarred  off  into  a 
car.  The  ore  is  fed  to  the  machine  through  a  pug-mill  where  it  is  wet  to 
4  per  cent  moisture  by  a  spray,  and  falls  to  a  feed-hopper  set  4  in.  above  the 
pallets,  thus  ensuring  a  layer  of  ore  of  that  depth.  It  moves  along  at  the 
rate  of  12  to  20  in.  per  minute.  As  it  reaches  the  ignition-box  it  is  set 
afire  by  an  ignition-furnace  burning  coal,  or  otherwise  gasoline  is  burned, 
using  a  series  of  Bunsen  burners  supplied  by  air  under  pressure.  The 
suction-box  12  ft.  6  in.  long  by  30  in.  wide  is  connected  to  an  exhaust-fan  at 
a  vacuum  of  6  oz.  As  the  ore  passes  the  burner  it  is  ignited,  and  the  air, 
sucked  down  through  the  layer  of  ore,  continues  the  burning,  which  is  com- 
pleted by  the  time  the  ore  reaches  the  end  of  the  suction-box.  There 


<SIDE  ELEVATION 


.SECTION  THROUGH   MACHINE 


FIG.  83. — Straight-line  Dwight-Lloyd  Blast-roasting  Furnace. 

results  a  product  which  retains  3  to  5  per  cent  of  sulphur  only.  By  the 
tune  the  ore  layer  has  reached  the  discharge-end  it  is  solid.  The  mouth  of 
the  feed  hopper  is  set  to  give  the  required  depth  of  material,  and  the  layer 
of  ore  is  smoothed  by  a  stiff  brush  30  in.  wide,  then  by  a  roller,  all  t eliding 
to  level  and  compact  the  ore  to  ensure  even  sintering  and  roasting.  Care 
must  be  taken  that  the  layer  of  ore  is  of  uniform  density  and  that  there  is 
no  segregating  of  it  in  the  feed-hopper;  otherwise  the  coarser  part  of  the 
layer  burns  rapidly  while  the  denser  part  does  not  get  enough  air. 

The  mixture  or  charge  may  consist  of  lead  concentrate,  fine  oxidized 
and  silicious  ores,  together  with  flue  dust.  Its  composition  may  be  quite 
variable.  A  satisfactory  mixture  will  carry  35  per  cent  SiC>2,  18  per  cent  S, 
and  20  per  cent  Pb;  another  one,  16  per  cent  SiO2,  18  per  cent  Fe,  15  per 
cent  S,  and  30  per  cent  Pb. 


SINTER-ROASTING  REACTIONS  113 

The  machine  will  treat  40  tons  in  twenty-four  hours  at  a  cost  of  75 
cents  per  ton.  One  to  two  horse-power  is  needed  to  drive  it. 

Reactions  in  Sinter-roasting  of  Copper-bearing  Sulphides.  —  When  a 
mixture  of  iron  and  copper  sulphide  is  sinter-roasted,  desulphurization 
proceeds  rapidly  if  the  ore  be  wet  and  silica  be  added;  otherwise  it  pro- 
ceeds slowly.  For  these  reactions  we  have.: 


(15) 

(16)  2Fe2O3+7H2S=4FeS+3SO2+14H. 

When  air  is  drawn  through  the  charge  both  hydrogen  and  H2S  burn, 
but  FeaO4  reacting  on  FeS  gives  FeO  as  follows: 

(17)  FeS+3Fe3O4=  10FeO-hSO2. 

This  reaction  is  exothermic  and  at  a  high  temperature  with  silica  would 
form  ferrous  silicate,  again  producing  heat.  Indeed,  in  action  the  forma- 
tion of  this,  with  the  consequent  sintering,  can  be  seen  spreading  as  the 
burning  proceeds.  The  cost  of  blast-roasting  in  pots  has  been  given  at 
$1.44  to  $1.80  per  ton. 

TRIPLE  ROASTING 

While  blast-  or  pot-roasting  is  less  used  than  formerly,  in  one  case,  viz., 
at  East  Helena,  Mont.,  zincky  Coeur  d'Alene  concentrates  high  in  lead 
are  roasted  in  three  stages.  A  Godfrey  revolving-hearth  roaster  brings  it 
down  13  per  cent  sulphur.  Again  crushed,  it  passes  to  a  Dwight-Lloyd 
sinter  machine  which  reduces  the  sulphur  to  8  per  cent.  This  product,  if 
smelted,  would  produce  too  much  matter,  therefore,  after  again  crushing,  it 
is  blast-roasted  in  Huntington-Heberlein  pots,  and  yields  a  final  material 
that  carries  but  2  per  cent  of  sulphur  —  so  low  is  this  element  in  fact  that 
there  is  no  production  of  matte  in  smelting  it. 


CHAPTER  X 

CONCENTRATION  OF  ORES  AS  A  SUBSIDIARY  OPERATION  IN 

METALLURGY 

In  mills  treating  ore  containing  a  heavier  part,  such  as  sulphides, 
gold  or  amalgam  from  the  plates,  it  is  customary  to  catch  such  heavy 
products : 

CONCENTRATION 

1.  By  gravity  concentration  in  which  the  heavy  particles  of  the  pulp 
are  caught  on  blanketed  or  canvas-covered  surfaces; 

2.  By  concentrating  tables; 

3.  By  oil  notation. 

1.  The  Blanket  or  Canvas  Table. — This  consists  of  a  floor,  sloping 
like  an  amalgamating  plate,  1|  in.  to  3  in.  per  foot.  It  is  smoothly  covered 


FIG.  84.— \Vilfley  Concentrating  Table. 

with  blankets  or  canvas,  and  along  its  upper  edge  the  mill-pulp  is  flowed  in 
an  even,  thin  layer  from  a  head  launder  to  distribute  over  the  table.  The 
lighter  particles  flow  down  the  slope,  the  heavy  portion  sinking  into  the 
interstices  of  the  fabric.  After  a  while  the  mill-flow  is  switched  to  another 
table  or  floor.  Fresh  water  or  solution  is  then  run  on  the  loaded  table 
which  sweeps  away  the  settled  pulp,  leaving  the  heavy  particles  in  the 
interstices  of  the  blanket  or  canvas. 

114 


THE  WILFLEY  CONCENTRATOR 


115 


2.  The  Concentrating  Table. — There  are  different  varieties  of  these, 
such  as  the  Frue  Vanner  or  the  Wilfley  table.  The  mill-flow  is  received  in 
a  continuous  stream  at  the  head  launder  of  the  machine  and  delivers 
two  products,  a  head  product  or  concentrate,  and  a  tail  product  freed  from 
the  heavy  portion. 

The  Wilfley  Table. — This  may  be  taken  as  a  type  of  table,  which  like 
the  Overstrom,  the  Deister,  or  the  Burchart  is  much  in  use  in  mills.  Fig. 
84  is  a  front  view  of  a  No.  6  Wilfley  table  having  a  deck,  a  surface  partly 
riffled,  partly  smooth,  so  transversely  inclined  that  a  sheet  or  film  of  water, 


FIG.  85.— Wilfley  Table  in  Action. 


composed  of  feed-water  conveying  ore-pulp  and  the  washing  or  dressing 
water  may  be  caused  to  flow  across  it.  This  deck,  mounted  on  bearings, 
has  an  endwise  reciprocating  motion  of  J  to  f  in.,  imparted  by  a  mechanical 
device  called  the  "  head  motion."  This  head  motion  causes  a  separation 
of  the  heavy  and  light  grains  into  layers  by  its  agitation  and  jerking  action, 
throwing  them  toward  the  head  end,  the  lightest-grained  being  washed 
down  the  slope  and  dropping  over  the  edge  at  the  tailings  side.  This 
separation  is  well  shown  in  Fig.  85.  The  heavy  grams  travel  along  in  the 
riffles  and  are  forced  more  toward  the  head,  escaping  there  and  at  the  head- 
end-corner. Each  kind  is  caught  in  its  own  launder. 


116       CONCENTRATION  OF  ORES  AS  A  SUBSIDIARY  OPERATION 


OIL  FLOTATION 

Flotation  is  a  method  of  concentrating  whereby  a  finely  ground  water 
pulp,  containing  sulphides,  as  of  lead,  copper,  zinc,  or  magnetite,  together 
with  a  gangue  of  quartz  or  other  equivalent  mineral,  is  treated  by  addition 
of  oil  (or  other  chemicals),  and  violently  agitated.  The  oil  or  other  chem- 
icals act  to  produce  a  film  upon  the  sulphide  particles,  causing  them  to 
float  as  a  froth  upon  the  surface  of  the  water,  while  the  gangue  sinks  to  the 
bottom.  The  froth  is  removed,  settled,  and  filtered,  yielding  a  product 
containing  the  valuable  sulphides.  This  method  possesses  the  advantage 
that  it  is  effective  upon  slimes  hard  or  impossible  to  concentrate  by  gravity 
methods. 

Most  plants  use  a  mixture  of  various  oils  and  the  kind  and  quantity 
should  be  worked  out  for  each  particular  case.  Pine  oil  is  quite  commonly 


ELEVATION  CROSS-SECTION 

FIG.  86. — The  Minerals  Separation  Machine.  . 

used,  also  wood  and  coal-tar  creosote.  Creosol  or  cresylic  acid  is  used  in 
small  proportions  with  coal-tar.  Other  substances  are  fuel-oil,  oleic  acid, 
and  eucalyptus  oil. 

Minerals  Separation  Flotation  Machine.— Referring  to  the  two  views  of 
a  double  machine,  Fig.  86,  the  feed  enters  the  first  agitation  box  at  the 
motor  end  of  the  machine,  thence  it  passes  to  a  second  box,  through  an 
opening  in  the  partition  wall,  as  shown  in  the  transverse  section. 

From  the  second  box  it  descends  to  the  first  spitzkasten  where  the  {roth, 
which  rises  to  the  surface,  flows  over  the  front  edge  into  a  narrow  deep 
launder.  The  remaining  pulp  passes  through  a  pipe  (the  inlet  controlled 
by  a  valve)  to  the  third  agitating  box.  From  this  box  the  pulp  passes  to 
the  second  spitzkasten,  and  so  on  through  the  machine  until  it  reaches  the 
fourteenth.  The  discharge  from  the  fourteenth  spitzkasten  leaves  the 
machine  as  tailings.  The  catch-launder  is  so  divided  that  the  first  four 
to  seven  spitskastens  make  frothed  concentrates,  the  remainder  make 
middlings  that  are  returned  to  the  system.  Some  of  the  pulp  is  over- 
flowed from  the  last  three  together  with  the  froth.  For  affecting  the  agi- 


FLOTATION  MACHINES  117 

tation,  each  of  the  sixteen  vertical  shafts  running  at  250  R.P.M.  carries 
a  four-bladed  impeller.  At  Anaconda  6  to  8  Ib.  of  60  per  cent  sulphuric 
acid,  2  to  3  Ib.  of  kerosene  sludge  acid,  and  i  to  1  Ib.  of  wood  creosote  is 
used  per  ton  of  flotation  feed.  Part  of  the  creosote  is  added  ahead  of 
the  tube-mill,  where  the  pulp  is  ground,  the  remainder  with  the  acids  at 
the  first  agitation  box.  The  pulp  is  heated  to  70°  F.  by  blowing  live 
steam  into  it  at  the  head  of  the  machine.  A  machine  will  treat  175 
tons  daily  of  slimed  pulp.  Other  flotation  machines  much  used  are  the 
Callow  and  the  Janney. 


PART  II 
GOLD 


CHAPTER  XI 

GOLD  ORES  AND  CLASSIFICATION  FOR  MILLING 
OCCURRENCE 

Gold  occurs  in  nature,  both  in  the  native  state  and  combined  with  tel- 
lurium. 

Native  gold  occurs  in  vein-matter  disseminated  in  grains  or  particles  of 
various  sizes,  and  it  is  found  not  only  in  quartz  veins,  but  in  veins  or  lodes 
containing  hematite,  iron-pyrite,  arsenical-pyrite,  blende,  and  galena. 
In  pyrite  it  occurs  not  only  in  the  substance  of  the  crystals,  but  as  films  on 
the  surface  of  these  crystals.  It  is  frequently  accompanied  by  silver. 
When  gold-bearing  veins  have  become  disintegrated  and  swept  away  into 
alluvial  deposits,  the  particles  of  gold,  where  released,  are  found  in  the 
sand  and  gravel  of  the  beds,  the  pebbles  and  boulders  themselves  (which 
have  come  from  the  country  rock),  being  in  general  barren  of  gold.  Gold 
occurring  in  this  way  is  called  alluvial  gold,  and  is  recovered  by  methods 
of  hydraulic  mining  or  dredging,  which  belong  to  mining  engineering 
rather  than  to  metallurgy.  We  shall  consider,  therefore,  the  treatment 
of  gold  ore. 

Gold  Tellurides. — In  South  Dakota,  at  Cripple  Creek,  Colo.,  in  West- 
ern Australia  and  elsewhere  is  to  be  found  gold  combined  with  tellurium 
as  calaverite,  AuTe2  (containing  41.4  per  cent  Au  and  57.3  per  cent  Te); 
also  gold  and  silver  combined  with  tellurium  as  sylvanite  (AuAg)Te2,  and 
as  petzite  (Ag2Te,  Au2Te). 

Physical  Properties  of  Gold. — This,  the  only  yellow  metal,  has  a  specific 
gravity  of  19.3  and  is  the  most  malleable  and  ductile  of  all  metals  so  that 
an  ounce  of  it  will  cover  160  sq.  ft.  It  is  softer  than  silver,  harder  than 
tin,  and  has  a  tenacity  of  14,000  Ib.  per  square  inch,  with  a  30.8  per  cent 
elongation.  It  melts  at  1063°  C.  and  begins  to  volatilize  at  1100°  C.,  so 
that  by  the  time  it  reaches  1250°  C.  it  volatilizes  four  times  as  fast. 

These  metals  are  commonly  found  together  in  ores,  and  a  metallurgical 
method  suited  for  the  extraction  of  one  is  often  as  well  suited  to  the  recovery 
of  the  other.  The  recovered  metals,  alloyed  with  one  another,  and  forming 
a  gold-silver  bar,  are  shipped  to  the  mint  or  to  the  refiner  in  that  form  for 
final  parting  into  their  constituent  metals.  Both  gold  and  silver  are  won 
from  their  ores  by  milling  or  by  smelting  methods,  the  former  being  the 
commoner  way. 

121 


122  GOLD  ORES  AND  CLASSIFICATION  FOR  MILLING 

Valuation  of  Gold  and  Silver. — The  amount  of  gold  and  silver 
in  ores  is  expressed  in  ounces,  in  pennyweights,  in  grams  or  in  kilo- 
grams per  ton.  When  in  the  form  of  bars  or  ingots  the  value  is 
given  in  the  percentage  or  in  the  fineness  of  the  respective  metals. 
In  English-speaking  America  silver  and  gold  in  ores  and  by-products 
are  designated  in  ounces  per  ton  or  in  dollars.  Thus  0.05  oz.  equals 
$1  per  ton,  valuing  gold  at  $20  per  ounce.  Its  exact  mint  value  is, 
however,  $20.67  per  ounce.  In  the  British  Empire,  except  Canada,  notably 
in  South  Africa  and  Australia,  the  pennyweight  is  preferred,  a  convenient 
designation,  since  it  gives  the  value  very  close  to  $1  or  four  shillings  English 
money.  In  Latin  America  gold  is  expressed  in  grams,  silver  in  kilograms 
per  metric  ton.  The  contents  of  a  gold-silver  bar  is  given  in  fineness,  i.e., 
in  parts  per  thousand.  Thus  such  a  bar,  upon  assay,  may  be  750  fine 
in  gold,  150  fine  in  silver,  the  remaining  100  parts  being  mainly  copper. 
United  States  gold  and  silver  coins  are  900  fine  in  gold  and  silver  respect- 
ively, the  100  parts  remaining  being  copper.  Sterling  gold  or  silver  is 
925  fine. 


CLASSIFICATION  FOR  MILLING 


Referring  to  the  milling  of  gold  ores,  that  is,  to  ores  in  which  gold  is  the 
dominant  metal,  these  may  be  treated  directly,  or  as  supplemented  by  con- 
centration as  follows : 
^  Amalgamation; 

Chlorination ; 
^  Cyaniding; 

Amalgamation  and  cyaniding; 

Amalgamation,  concentration,  and  cyaniding  the  tailings; 

Concentration  and  cyaniding  the  tailings. 

Where  concentrates  are  made  these  are  further  treated  by  smelting,  01 
at  the  mill  by  cyaniding  them.  They  may  be  produced  either  by  gravity 
concentration  or  by  flotation. 


CHAPTER  XII 
AMALGAMATION 

Plate  Amalgamation. — This  is  the  time-honored  method  for  the  recovery 
of  gold  from  "  free-milling  "  ores  and  as  an  important  preliminary  process 
with  gold  ores,  which  though  not  free-milling,  contain  a  considerable  pro- 
portion of  readily  amalgamable  gold.  By  free-milling  ores  we  mean  those 
in  which  the  gold  occurs  native  and  can  be  caught  by  amalgamation. 
However,  the  cyanide  process  is  used  after  amalgamation  where  a  portion 
of  the  gold  is  not  recoverable  by  amalgamation. 

Ores  Needing  Amalgamation. — Free-milling  oxidized  gold  ores  in  which 
the  gold  is  shown  upon  panning,  and  where  many  of  the  particles  are 
too  coarse  for  solution  by  cyaniding,  are  the  ones  for  this  method.  At  the 
Manhattan  Big  Four  mill,  where  the  ore  is  a  soft  calcareous  schist  with 
laminations  containing  calcite  and  quartz,  but  no  sulphides,  the  fine  gold, 
occurring  in  the  laminations,  is  stamp-  and  tube-milled  as  just  above  out- 
lined. It  is  thus  seen  that  when  the  ore  has  been  stamped  coarse,  the  pulp 
is  not  run  over  the  amalgamating  plates  until  finely  ground,  so  hi  this  mill 
we  find  the  plates  not  in  their  customary  place  below  the  mortar,  but  in  a 
separate  building,  where  they  receive  special  attention,  and  are  safe  from 
pilfering. 

Whether  inside  amalgamation  is  the  most  suitable  is  often  discussed; 
but  for  certain  ores  it  is  the  correct  method. 

Gold  sometimes  occurs  in  certain  ores  with  a  metallic  appearance,  but 
is  mostly  brown  and  lusterless.  In  this  supposedly  allotropic  form  it 
fails  to  attach  itself  to  the  amalgamated  plate. 

There  is  a  considerable  difference  of  opinion  among  metallurgists  regard- 
ing the  extent  to  which  amalgamation  should  be  used  in  a  goldmill. 
Some  think  with  very  fine  grinding  to  bring  all  the  gold  (and  any  silver) 
to  the  degree  of  subdivision  necessary  for  rapid  solution,  that  amalgamation 
may  be  dispensed  with,  and  that  by  "  all-sliming  "  the  gold  can  be  extracted 
by  cyaniding. 

The  two  extremes  in  the  plate  area  provided  for  amalgamation  prior  to 
cyanidation  are  represented  by  the  Rand  and  the  Homestake.  Figures 
from  thirteen  Rand  plants,  for  the  year  1913,  show  the  recovery  by  amalga- 
mation on  mill  feed,  approximating  $6,  to  be  61.4  per  cent  the  plate  area 
ranging  from  0.25  to  2.0  sq.  ft.  per  ton  milled  per  day,  and  the  mercury 

123 


124  AMALGAMATION 

consumption  about  0.1  oz.  per  ton  milled.  The  tendency  on  the  Rand  has 
been  to  eliminate  amalgamation  directly  after  the  stamps  in  favor  of 
amalgamation  after  tube-milling,  and  at  the  same  time,  reduce  the  plate 
area  to  less  than  a  half.  In  no  case  has  the  reduced  plate  area  caused  serious 
decrease  in  recovery  by  amalgamation,  some  instances  showing  no  decrease, 
while,  naturally,  there  is  a  decided  saving  in  mercury-loss,  attendance,  and 
capital  cost.  On  the  other  hand,  amalgamation  at  Homestake  is  carried 
on  both  inside  and  outside  of  the  stamp  mortars  and  after  regrinding.  Of 
the  total  plate  area  of  approximately  11  sq.  ft.  per  ton  milled,  only  1.60 
per  cent  is  used  after  regrinding.  In  1914,  on  $4.11  mill  feed,  the  Home- 
stake  recovery  by  amalgamation  at  the  stamps  was  69.0  per  cent,  and  after 
regrinding,  0.8  per  cent;  the  total  of  69.8  per  cent  being  effected  with  a  loss 
of  mercury  of  0.13  oz.,  and  at  a  total  cost  of  2.45  cents  per  ton  milled. 

Amalgamation,  as  practiced  at  the  average  plant,  costs  between  5  and 
10  cents,  the  plate  area  approximating  2  sq.  ft.  per  ton  milled.  The 
recent  trend  has  been  to  relegate  to  this  step  the  recovery  of  such  coarse 
gold  as  may  greatly  retard,  or  entirely  escape  subsequent  cyanide  treat- 
ment. With  amalgamation,  crushing  is  nearly  always  done  in  water, 
although  a  few  plants  have  been  able  to  maintain  their  plates  in  fair  con- 
dition when  crushing  in  cyanide  solution.  The  extremely  low  cost  of 
recovering  gold  by  amalgamation  demands  that  this  step  be  given  weighty 
consideration.  An  unusual  method  of  amalgamation  at  the  Nipissing 
high-grade  mill  is  to  be  noted.  A  charge  of  3  tons  of  2500-oz.  silver  ore 
(containing  39  per  cent  arsenic,  9  per  cent  cot>alt,  6  per  cent  nickel),  4  tons 
of  mercury  and  1.5  tons  of  KCN  solution  is  tube-milled,  with  compressed 
air  fed  into  the  tube-mill  for  some  ten  hours,  at  the  end  of  which  period 
amalgamation  has  recovered  some  97  per  cent  of  the  silver,  the  residues, 
freed  from  amalgam  and  mercury,  then  going  to  the  cyanide  plant.  The 
mercury  consumption  is  20  Ib.  per  ton  of  ore. 

STAMP-MILLING  WITH  PLATE  AMALGAMATION 

The  ore  is  crushed  to  a  size  of  20-mesh  or  finer,  using  6  to  8  tons  of  water 
per  ton  of  ore,  and  running  the  ore  pulp  over  plates  of  at  least  4J  ft.  wide 
by  6  ft.  long  per  battery  or  often  much  longer.  The  heavier  gol<^  soon 
touches  the  plate  where  it  becomes  incorporated  with  the  amalgamated 
surface.  About  every  shift  the  battery  is  stopped  for  a  few  minutes  and 
the  gold-bearing  amalgam  is  scraped  from  the  plates,  and  treated  to  obtain 
the  gold.  Amalgamated  plates  may  also  be  placed  inside  the .,- mortar; 
particles  of  gold  adhere  to  these  plates  when  driven  against  them  by  the 
splash  of  the  pulp.  Gold  particles  also  fall  to  the  bottom  of  the  mortar  to 
be  caught  by  mercury  there.  The  gold  caught  inside  the  mortar  is  recovered 
in  the  monthly  clean-up.  From  time  to  time  about  1.5  oz.  mercury  per 
ounce  of  gold  in  the  ore  is  added  in  the  mortar  as  the  crushing  proceeds. 


THE  STAMP  BATTERY 


125 


Plate  amalgamation  is  used  for  gold  ores  only. 

The  tailing,  if  barren,  is  run  to  waste.  If  it  contains  gold-bearing  pyrite 
this  may  be  caught  on  concentrating  tables,  see  Fig.  84,  to  be  sent  away 
for  smelting  or  to  be  specially  treated  on  the  spot  by  cyaniding.  The 
tailing,  generally  gold-bearing,  would  be  cyanided. 

The  Stamp-battery. — Fig.  87  is  a  view  of  a  ten-stamp  battery  of  wood 


. — Perspective  View  of  Ten-stamp  Battery. 

construction.  The  parts  are  thus  designated.  At  A,  the  mortar  block  or 
foundation;  B,  the  mud  sills,  C,  the  cross  sills;  D,  the  side  posts;  F  and  G, 
the  buck-staves;  H  and  7,  the  lower  and  upper  guide-timbers  respectively. 
The  foregoing  parts  constitute  the  battery  frame.  «/,  J  are  the  cast-iron 
mortars  as  shown  in  section  in  Fig.  88.  At  K  is  a  wire  cloth,  or,  as  here 
shown,  a  slotted  screen.  At  L  is  the  die,  resting  on  the  sole  of  the  mortar. 
The  stamp  consists  of  a  stem  having  on  it  tappet  P,  by  which  it  is  lifted  and 


126  AMALGAMATION 

at  its  lower  end  a  boss  N,  carrying  its  shoe  M.  these  four  parts  being  the 
total  weight.  In  Fig.  88  this  weight  is  1250  Ib.  R  is  the  cam  shaft  carry- 
ing ten  cams  U  (better  shown  in  Fig.  88),  with  its  driving  pulley  V  at 
one  end.  At  Y  is  seen  one  of  the  amalgamated  plates,  already  referred  to, 
the  other  being  omitted  to  show  the  screen  and  chuck-block  belonging 
to  the  first  five  stamps. 

In  Fig.  87  the  mortar  blocks  are  heavy  posts,  set  on  end,  and  extending 
down  to  the  solid  rock  or  concrete  foundation.  Instead  of  these,  concrete 
blocks,  Fig.  88,  are  preferred,  the  mortars  being  held  down  by  long  foun- 
dation bolts  with  j-in.  sheet  rubber  and  f  in.  sheet-lead  interposed  to 
give  an  even  bearing. 

Fig.  88  is  a  sectional  elevation  of  a  ten-stamp  battery  unit,  grouped  as 
shown  in  Fig.  87,  into  two  sets  of  five  stamps,  Ore  from  the  feed  bin  is  fed 
to  the  stamps  by  a  suspended  Challenge  ore-feeder,  which  can  be  run 
out  of  the  way  when  making  repairs  to  the  stamps.  At  S  is  shown  a  hori- 
zontal lever  depressed  at  each  stroke  by  a  collar  on  one  of  the  stamps  and 
by  a  vertical  link  and  lever  working  a  ratchet  feed.  As  the  circular  feed- 
plate  slowly  revolves  against  a  fixed  scraper,  the  ore  is  scraped  off  to  fall  into 
the  mortar.  As  the  ore  accumulates  under  the  stamps,  the  stroke  shortens 
and  with  it  the  feed,  while  as  the  mortar  empties  the  stroke  is  increasesd 
and  with  it  the  feed.  The  ore  from  the  lip  of  the  feeder  falls  into  the 
mortar,  there  to  be  stamped,  and  when  sufficiently  fine,  the  resultant  ore 
pulp  is  driven  through  the  screen  at  the  front  by  the  splash  caused  by  the 
dropping  stamps,  to  flow  over  the  apron  plates  (not  here  shown).  The 
guide  timbers  carry  the  lower  and  upper  cast-iron  guides  for  each  stamp 
stem.  At  X  is  a  finger-bar,  one  to  each  stamp,  by  which  the  stamps  are 
"  hung  up  "  when,  for  any  reason,  a  battery  is  to  be  stopped  without  stop- 
ping the  other  battery. 

Operation  of  the  Stamp-battery. — The  feed  is  regulated  so  as  to  cause 
the  stamps  to  strike  with  a  sharp,  hard  blow,  but  with  little  of  the  rebound 
that  would  occur  with  a  thin  layer  of  ore. 

Mercury  Fed  to  the  Battery. — This  will  average  1.5  oz.  per  ounce  of 
gold  caught.  Added,  a  little  at  a  time  inside  the  mortar,  it  works  out 
in  part  upon  the  apron  plates.  For  the  amount  to  be  used  the  mill-man  is 
guided  by  the  appearance  of  the  plates.  If  they  are  hard  it  means  too  little 
mercury  or  "  quick  ";  if  the  mercury  shows  on  them  in  streaks  or  patches 
then  too  much  is  being  fed.  Mercury  should  be  free  from  base  metals 
that  would  cause  it  to  "  sicken  "  into  coated  globules,  which  would  be 
swept  away  with  the  pulp.  It  is  better  that  it  contain  a  little  gold.  * 

*  The  loss  of  mercury  may  aggregate  0.5  oz.  per  ton  of  ore  treated.  It  may  be  lost 
by  flouring,  indicated  by  a  white  appearance,  and  due  to  excessive  agitation  in  the  air, 
which  breaks  it  into  particles  so  fine  that  they  unite  no  more.  Mercury  may  be  lost 
by  "  sickening/'  shown  by  a  black  appearance  and  due  to  the  presence  of  base  metals 
as  already  explained. 


THE  STAMP  BATTERY 


127 


FIG.  88. — Sectional  Elevation  of  Ten-stamp  Battery. 


128 


AMALGAMATION 


Dressing  the  Plates. — This  is  done  three  or  four  times  daily,  and  takes 
about  fifteen  minutes.  To  do  this  feeding  is  stopped  so  that  the  ore  may 
work  out  of  the  mortar,  and  the  stamps  are  hung  up.  The  amalgam  on 
the  plates,  perhaps  a  cupful,  is  removed  with  a  rubber-edged  scraper.  If 
the  surface  of  the  plate  is  hard  a  little  mercury  is  sprinkled  on  it. 
Where  the  plate  has  become  tarnished  by  a  verdigris  coating  this  may 
be  removed  by  salammoniac  applied  with  a  scrubbing  brush.  In  a  few 


FIG.  89. — Clean-up  Pan. 


I 


minutes  this  should  be  washed  off,  and  potassium  cyanide,  then  mercury 
rubbed  on,  and  the  plate  washed  clean.  The  stamps  are  now  started,  and 
feeding  is  resumed. 

Apron-plates  are  set  at  a  grade  of  J  to  If  in.  per  foot,  the  steepest  grade 
where  sulphides  occur  in  the  pulp.  When  properly  flowing  over  the  plate, 
the  pulp  travels  down  in  a  series  of  ripples,  thus  bringing  the  gold  in  con- 
tact with  it. 

A  mercury-trap  receives  the  flowing  pulp  at  the  foot  of  the  plate. 


THE  CLEAN-UP  PAN 


129 


Any  non-adherent  particles  of  amalgam  are  here  caught.  The  amalgam 
is  occasionally  removed  by  the  plug-hole  on  the  bottom  of  the  trap.  Its 
overflow  passes  generally  to  the  concentrating  tables. 

The  Clean-up. — This  occurs  once  or  twice  a  month.  Let  us  take  the 
case  of  a  40-stamp  mill.  Two  batteries  are  hung  up.  The  screens,  inside 
plates,  and  dies  are  taken  out.  The  ore  in  the  mortars,  two  or  three 
bucketsful,  are  taken  and  fed  to  the  next  batteries.  The  inside  plates  are 
scraped  and  dressed,  and  all  is  replaced,  and  the  batteries  again  started. 
The  next  two  are  treated  in  the  same  way  as  well  as  the  last  ones,  whose 
mortar  contains  an  accumulation  from  them  all. 

The  Clean-up  Pan,  Fig.  89,  3  ft.  diameter  and  making  12  to  15  R.P.M., 


FIG.  90—  Vertical  Retort. 

is  used  for  grinding  the  sand,  pyrite,  fragments  of  iron,  etc.,  the  accumula- 
tion of  the  last  batteries.  The  charge,  perhaps  300  lb.,  is  wet-ground  to  a 
fine  mud  with  the  addition  of  50  lb.  of  mercury  during  the  shift.  The 
pulp  is  diluted  with  water  and  the  muddy  portion  is  run  off  by  a  side  plug. 
The  residual  mercury  and  amalgam  with  some  mud  is  withdrawn  through 
the  lowest  plug-hole,  panned  in  a  gold-pan  by  hand,  treated  with  nitric  acid, 
and  well  washed  until  clean.  The  residual  amalgam  is  strained  through 
canvas  to  remove  the  excess  mercury.  Gold  amalgam  thus  treated  con- 
tains 35  to  45  per  cent  gold;  the  filtered  mercury  still  retains  0.5  per  cent 
gold. 

Retorting. — In  the  smaller  gold  mills  the  amalgam  is  retorted  as  shown  in 
Fig.  90.     The  retort,  filled  two-thirds  full  of  amalgam  and  the  cover  luted 


130  AMALGAMATION 

and  clamped,  is  placed  in  a  wind-furnace,  there  supported  on  a  cast-iron 
rest.  A  water-cooled  pipe  leads  out  from  the  cover,  its  lower  end  dipping 
in  a  tub  of  water.  The  retort  is  gradually  heated,  the  mercury  vapor 
comes  over,  and  is  condensed  in  drops  in  the  water-cooled  pipe  and  collects 
in  the  tub  below.  The  retort  is  kept  at  a  distilling  temperature  for  one 
or  two  hours,  then  heated  to  redness  to  expel  the  last  of  the  mercury. 
This  mercury  is  used  again. 

The  residue,  taken  from  the  retort,  is  porous  and  is  from  500  to  900  fine 
in  gold.  It  is  melted  in  a  wind-furnace,  see  Fig.  64,  with  soda  and  borax, 
and  when  it  contains  base  metal,  with  an  addition  of  a  little  niter,  which 
serves  to  toughen  it.  The  melt  is  poured  into  an  ingot-mold  and,  after 
cooling,  cleaned  from  adhering  slag  and  shipped  to  the  mint. 

GENERAL  ARRANGEMENT  OF  A  GOLD  STAMP-MILL 

Fig.  91  indicates  clearly  the  course  of  the  ore  through  the  mill,  while 
Fig.  139,  taken  to  the  end  of  the  concentrating  tables,  gives  a  general  idea 
of  such  a  mill. 

Run-of-mine  ore  enters  the  mill  at  the  highest  level  and  is  dumped 
into  the  storage  bin  A,  thence  it  is  withdrawn  through  a  Blake  ore-breaker 
6,  which  discharges  to  the  feed-bin  C.  The  feed-bin  is  filled  during  the 
day-shift,  and  is  large  enough  to  hold  a  twenty-four-hour  supply.  From 
the  bin  the  ore  is  drawn  off  through  a  regulated  sliding  gate  by  chute  to  the 
suspended  Challenge  ore-feeder,  and  thus  is  in  constant  supply  to  the 
stamp  battery,  where  it  is  crushed  with  an  addition  of  mercury  and  of 
water.  The  pulp,  splashing  through  the  battery  screens,  flows  over  amal- 
gamated plates  e,  where  the  gold  is  caught.  The  tailing  from  the  plates 
unites  in  a  launder  and  finally  falls  into  a  distributing  box  that  commands 
four  concentrating  tables.  A  distribution  is  made  here  and  one-fourth  the 
flow  is  supplied  by  a  launder  to  each  table.  The  tailing  from  the  tables  is 
wasted.  The  concentrate  is  collected  and  shipped  for  smelting.  The 
method  of  driving  the  machines  is  indicated  in  Fig.  88.  Above  the  battery 
runs  an  overhead  track  carrying  a  trolley  and  a  heavy  chain  tackle  by  means 
of  which  parts  can  be  removed  readily  or  replaced.  ^ 

In  practice  stamps  vary  from  as  light  as  850  Ib.  to  as  high  as  2000  Ib. 
as  on  the  Rand,  South  Africa.  Many  stamps  in  American  practice  are  of 
1000  Ib.  though  there  are  mills  having  them  of  1250  Ib.  and  1500  Ib.  The 
tendency  at  present  is  toward  superseding  them  by  rolls  or  by  ball-mills,  or 
supplementing  them  by  tube-mills.  In  such  cases  coarser  screens  are  put 
in  at  the  stamps,  their  duty  is  increased,  and  added  work  is  put  upon  the 
tube-mills,  which  then  grind  in  closed  circuit  with  Dorr  or  Akins  classifiers 
(see  Figs.  29  and  30). 


THE  STAMP-MILL  WITH  CONCENTRATION 


131 


CONCENTRATION  IN  STAMP-MILLING 

Methods. — Three  general  methods  of  concentration  are  practiced: 
(a)  removal  of  high-grade  concentrates  for  shipment  to  smelters;  (6) 
removal  of  lower-grade  concentrates  for  local  treatment  by  cyanidation,  or 


FIG.  91. — Stamp-mill  followed  by  Amalgamation  and  Concentration. 

roasting  followed   by  cyanidation ;  (c)  removal  of  concentrates  for  finer 
grinding  and  returning  to  the  regular  pulp. 

Systems. — A  very  complete  concentration  system,  on  a  complex  ore, 
is  carried  out  by  the  Goldfield  Consolidated,  where  approximately  6 
per  cent  by  weight  and  67  per  cent  by  value  of  the  feed  is  removed  at  a 
cost  for  concentration  of  6  cents  per  ton  of  ore  concentrated,  the  concen- 
trates then  receiving  a  very  successful  local  treatment.  On  the  26-oz. 
cilver  ore  of  the  San  Rafael  (Pachuca),  1J  per  cent  by  weight  and  22  per 


132  AMALGAMATION 

cent  by  value  is  taken  out  as  concentrate,  to  be  shipped  to  smelter.  On  the 
high-grade  Esperanza  (El  Oro)  sulphide  ores,  concentration  removed  1.2 
per  cent  by  weight,  with  35  per  cent  of  the  gold  and  16  per  cent  of  the  silver. 
Tonopah  concentration  removes  approximately  15  per  cent  of  the  silver; 
Stratton's  Independence  removes  approximately  44  per  cent  of  the  gold; 
Liberty  Bell,  8  per  cent  of  the  gold  and  20  per  cent  of  the  silver. 

Costs. — The  cost  of  concentration,  per  ton  of  ore  concentrated,  ranges 
from  5  to  15  cents — averaging  perhaps  10  cents  per  ton.  The  apparent 
saving  in  the  removal  of  refractory  value  and  cyanides  by  concentration, 
is  reflected  by  decreased  cyanide  consumption  and  solution  contact,  and 
by  lower  assay  value  of  final  plant  residue.  This  must  be  carefully  bal- 
anced against  the  higher  cost  of  realization  and  mechanical  loss  in  ship- 
ment of  concentrates. 


CHAPTER  XIII 

THE  HYDROMETALLURGY  OF  GOLD  ORES 
MILLING  ORES  IN  AQUEOUS  SOLUTIONS 

At  the  present  time  there  are  two  methods  by  which  gold  is  dis- 
solved from  its  ore  by  chemical  solvents.  In  either  process  the  first 
step  is  to  obtain  the  .gold  in  aqueous  solution,  then  to  precipitate  it 
from  the  clear  filtrate,  and  finally  to  get  it  in  the  form  of  a  bar  or  ingot. 

The  two  processes  are: 

(1)  The  chlorination  or  Plattner  process,  by  which  the  gold  is  obtained 
in  solution  as  a  chloride  by  the  action  of  an  aqueous  solution  of  chlorine 
gas. 

(2)  The  Cyanide  or  Mac  Arthur-Forrest  process,  in  which  the  solution  of 
the  gold  is  effected  by  a  weak  cyanide  solution,  the  dissolved  gold  then 
being  present  as  potassium  auro-cyanide.     With  certain  refractory  ores, 
the  activity  of  the  solution  is  greatly  increased  by  the  use  of  bromine  or 
bromo-cyanogen  in  addition  to  the  potassium  cyanide. 

Extraction  of  gold  by  means  of  a  solvent  in  aqueous  solution  is  also 
practiced  where  gold  cannot  be  completely  extracted  by  amalgamation. 
This  often  is  the  case  with  pyrite  ores;  and  extraction  can  be  practiced  to 
advantage,  not  only  where  amalgamation  is  unsuitable,  but  where  smelting 
is  expensive. 

Gold  in  ore  occurs  hi  particles  of  various  sizes,  both  as  grains  readily 
seen,  and  in  particles  of  microscopic  size.  When  the  particles  are  visible, 
or  when  the  ore  shows  "  colors  "  upon  panning,  the  gold  is  called  coarse, 
and  such  particles  generally  can  be  recovered  by  amalgamation.  Gold 
often  occurs  in  finely  disseminated,  microscopic  particles,  not  visible  to 
the  eye,  and  in  films  on  the  surface  of  pyrite  crystals.  If  the  ore  can  be 
ground  so  fine  as  to  unlock  the  crystals,  or  if  it  is  permeable  to  solutions, 
gold  can  be  dissolved  in  aqueous  solvents,  such  as  chlorine  or  potassium 
cyanide.  Advantage  is  taken  of  the  solubility  of  the  released  gold  particles, 
and  leaching  or  percolation  methods,  in  tanks  or  vats,  are  practiced  with 
this  in  view.  The  solution  soaks  through  the  ore,  comes  hi  contact  with 
gold  particles,  and  dissolves  them,  or  by  another  process,  the  finely  ground 
ore  or  slime  is  agitated  with  the  solution,  and  the  pulp  is  filtered  and 
washed  hi  filter-presses.  The  clear  filtrate,  in  any  case,  is  treated  by  a 

133 


134  THE  HYDROMETALLURGY  OF  GOLD  ORES 

suitable  precipitant  to  obtain  the  gold  in  small  bulk,  and  the  precipitated 
gold  is  melted  and  cast  in  the  form  of  a  bar  or  ingot  for  sale. 

There  are  three  stages  in  any  method  of  extracting  gold  by  aqueous 
solvents:  (1)  The  ore  is  finely  ground  and  when  refractory,  roasted,  to  con- 
vert the  gold  into  a  soluble  form,  and  render  it  accessible  to  the  solution. 
(2)  The  gold  is  extracted  from  the  ore  by  means  of  a  dilute  solvent,  using 
a  tank  with  a  filter-bottom,  or  agitating  the  ore,  pulverized  to  a  thin  pulp, 
using  a  filter-press  for  the  separation  of  the  solution.  (3)  The  gold  in  the 
solution  is  precipitated  (a),  in  chlorination  by  hydrogen  sulphide  or  other 
precipitating  agent  or  (6),  in  cyanidation  by  the  use  of  zinc-shaving  or 
zinc-dust.  The  precipitate  is  collected,  dried,  and  melted  into  an  ingot. 

Cyanidation  has  proved  to  be  a  remarkably  cheap  and  efficient  method 
of  extraction,  but  it  has  limitations,  not  only  in  respect  to  the  solubility 
of  the  gold,  but  because  of  the  interference  of  compounds  that 
sometimes  are  present,  notably  those  of  copper,  that  interfere  with 
extraction  in  various  ways.  The  process  has  the  advantage  over  the  chlo- 
rination method  in  that  silver,  as  well  as  gold,  can  be  extracted.  Under 
favorable  conditions  the  extraction  is  high,  and  modern  methods  have 
reduced  the  cost  of  treatment  to  a  low  figure. 

Pyrite  ore,  exposed  to  the  weather,  becomes  acid  in  reaction,  and  if 
treated  by  cyanide,  decomposes  and  destroys  the  potassium  cyanide. 
To  correct  this,  the  ore  is  first  treated  by  a  wash  of  dilute  caustic 
soda  or  of  acid  mixed  with  caustic  lime  in  sufficient  quantity  to  overcome 
acidity,  or  to  create  "  protective  alkalinity." 

When  ore  is  refractory  and  requires  preliminary  roasting,  this  adds 
much  to  the  cost  of  treatment.  In  chlorination,  roasting  is  always  neces- 
sary, and  in  any  case  it  improves  the  condition  of  the  ore  and  makes  it 
porous  and  permeable  when  leached  or  filter-pressed. 


CHAPTER   XIV 
(1)  CHLORINATION  OF  GOLD  ORES 

This  consists  in  attacking  the  gold  in  the  roasted  ore  with  chlorine  to 
form  the  soluble  gold  chloride,  and  dissolving  out  the  gold  chloride  in  water. 

ORES  SUITED  TO  CHLORINATION 

An  ideal  ore  for  chlorination  is  one  in  which  the  gold  is  present  in  a 
fine  state  of  division,  in  which  bases  are  absent  that  would  be  attacked  by 
chlorine,  and  silver  if  present  in  such  a  condition  as  not  to  coat  the  particles 
of  gold  with  insoluble  silver  chloride.  While  the  cyanide  process  is  better 
for  the  treatment  of  low-grade  ores,  many  refractory  high-grade  ores  have 
given  better  results  by  chlorination. 

Ores  in  which  the  gangue  consists  of  hydrated  iron-oxide  are  extremely 
difficult  to  amalgamate.  Not  only  is  the  gold  finely  divided,  but  the  ore 
is  slimy  and  forms  a  coating  on  the  amalgamating-plates.  Such  ores  give 
satisfactory  results  by  barrel  chlorination.  Silver  is  not  recovered  by  chlo- 
rination, since  it  becomes  an  insoluble  silver  chloride.  If,  however,  suf- 
ficient silver  be  present  to  pay  the  increased  cost,  salt  may  be  used  in  roast- 
ing and  the  silver  extracted  by  means  of  sodium  hyposulphite  or  better 
by  cyaniding. 

Ore  containing  sulphur,  arsenic,  and  antimony  is  crushed  to  10-  to 
30-mesh  size,  and  is  roasted  to  expel  these  elements,  to  oxidize  the  bases, 
to  leave  the  gold  in  such  form  as  to  be  attacked  by  chlorine,  and  to 
make  the  ore  porous,  accessible  to  chlorine,  and  more  easily  leached. 

Chlorination  of  Concentrate. — This  is  used  on  concentrate  from  gold- 
milling,  containing  much  sulphide,  and  typical  of  California  ore.  The 
coarse  gold  has  been  removed,  by  milling  and  amalgamation,  and  the 
concentrate,  generally  1.5  to  2  per  cent  of  the  weight  of  the  ore  milled, 
contains  gold  in  fine  particles.  It  is  roasted,  generally  in  a  long-bedded 
reverberatory  furnace,  see  Fig.  70,  60  ft.  long,  and  of  3  tons  capacity 
in  twenty-four  hours.  It  contains  copper,  lead,  lime,  and  magnesia, 
all  of  which  consume  chlorine,  and  form  chlorides.  To  prevent  this,  it  has 
been  customary  to  add  salt,  to  the  extent  of  0.75  to  1.5  per  cent  of  the  charge, 
at  or  near  the  completion  of  the  roast.  If  roasting  has  been  thorough  up 
to  this  time,  copper  is  present  as  CuO,  lead  as  PbSO4,  lime  as  CaO,  and 
magnesia  as  MgO.  Were  the  copper  present  at  the  end  as  CuSO4  it  would 

135 


136  CHLORINATION  OF  GOLD  ORES 

react  with  the  salt,  forming  a  chloride  of  copper.  The  common  salt  also 
reacts  upon  the  gold  and  forms  gold  chlorides.  Both  these  chlorides  are 
volatile,  and  the  CuCl,  in  volatilizing,  promotes  the  detrimental  volatiliza- 
tion of  the  gold. 

As  long  as  sulphur  is  present  it  protects  the  gold  from  attack,  but  when 
sulphates  have  been  formed,  and  are  causing  the  abundant  evolution  of 
chlorine  by  reaction  with  the  salt,  the  escaping  gas  carries  gold  chloride, 
the  gold  being  unprotected  by  sulphur  at  the  time  of  chlorination.  Lead 
sulphate  similarly  reacts  with  salt,  forming  lead  chloride,  which  does  not 
consume  chlorine.  When  much  lead  is  present,  however,  it  may  be 
removed  by  leaching  with  hot  water  before  treating  with  chlorine.  Lime 
and  magnesia  are  converted  by  the  salt  into  chlorides,  and  in  this  form 
consume  no  chlorine.  The  process  of  roasting  is  therefore  conducted  as 
follows: 

The  ore  is  thoroughly  roasted  at  a  low-red  heat.  The  temperature  is 
a  bright  red  (850°  C.),  to  decompose  copper  sulphate.  The  salt  is  added 
and  thoroughly  incorporated,  and  the  temperature  reduced  to  prevent 
volatilization  of  the  gold.  The  quantity  of  salt  to  be  added,  the  time 
needed  for  roasting,  and  the  temperature  compatible  with  the  minimum 
loss  of  the  gold,  should  be  determined  experimentally  for  each  kind  of  ore. 

THE  GOLDFIELD  CHLORINE  MILL  CO.,  GOLDFIELD,  NEV. 

At  this,  the  latest  development  of  vat-chlorination,  the  ore  of  the  100- 
ton  plant,  crushed  to  14-mesh,  receives  an  oxidizing  roast  in  a  muffle  type 
rabble-roaster,  is  cooled  and  moistened,  and  is  then  delivered  to  a  storage 
bin.  From  the  bin  it  is  charged  into  any  one  of  the  seven  wooden  leaching 
vats,  22  ft.  diameter  by  8  ft.  deep,  by  means  of  a  3-ton  grab-bucket  operated 
by  a  traveling  crane.  When  full,  a  wooden  cover  is  put  upon  the  vat,  and 
the  leaching  is  done  with  a  strong  solution  of  8  Ib.  of  chlorine  per  ton  of 
water.  The  filtrate  from  the  vat  is  pumped  to  a  solution-storage  tank 
and  is  thence  drawn  to  precipitating  boxes  (resembling  the  zinc  boxes  of  a 
cyanide  plant),  where  the  gold  is  precipitated  electrolytically.  In  these 
boxes  are  suspended  anode  plates  of  graphitized  carbon  and  cathod^  plates 
of  lead,  the  lead  having  been  alloyed  with  1  per  cent  of  zinc,  the  zinc  hasten- 
ing the  corrosive  action.  The  resultant  slime  of  the  electrolysis  consisting 
of  lead,  gold,  and  any  silver,  while  still  moist,  is  mixed  with  a  suitable 
flux  and  made  into  briquettes.  These  latter  are  melted  in  a  reverberatory 
furnace  as  rich  lead  bars,  are  refined  in  an  English  cupelling  furnace,  as 
described  under  "  Refining  Base  Bullion,"  page  498.  The  pumps  and 
piping  of  the  mill  are  made  of  rubber  to  withstand  the  corrosive  action  of 
the  chlorine. 

Making  the  Chlorine  Solution. — This  is  produced  electrolytically  from  a 


BARREL  CHLORINATION 


137 


brine  solution.  The  chlorine  ga«,  arising  from  the  electrolytic  tank, 
to  two  sheet-iron  stacks  lined  with  glazed  sewer  pipe  with  cement  between 
the  pipe  and  the  outer  iron.  Inside  these  is  placed  a  filling  of  perforated 
hollow  tile  balls.  Water  is  supplied  to  trickle  downward  through  the  balls, 
while  the  chlorine  gas  admitted  at  the  bottom  is  absorbed  by  the  water 
spread  upon  the  extended  surface  of  the  balls. 

BARREL  CHLORINATION 

This  process  was  evolved  as  being  better  suited  to  large  tonnages  of  ore 
than  the  vat  process,  and  its  flows-sheet  is  shown  in  Fig.  92. 

An  example  of  such  an  ore  is  that  of  Cripple  Creek.  The  ore  contains 
gold  telluride,  and  must  be  roasted  to  release  the  gold  from  combination 


Mine  Ore 


Crushing  Plant 

See  Fiff.  25 


Roasted  Cooled  Ore 


Crushed  jlaw  Ore 
I    Storage  Uins     | 

g         |  Belt  Kiev. -i tor  | 
|          I  Feed  Hopper  I 

£     |  Roasting  Fiirmtce~| 
Roasted  Ore 

I  Cooling  Hearth  &  Floor    | 


Solution 


/Precip-V-SOo  Gas 
I    Tank  J+  H2S  Gas 


Elevator 

* 
Storage  Bins 


Solution 
to  Waste 


Ingots 
to  Market 


Solution 


FIG.  92.— Flow-sheet  of  Barrel  Chlorination. 

with  tellurium,  and  to  expel  all  sulphur  above  0.1  per  cent.  The  com- 
plete process  of  barrel-chlorination  is  as  follows : 

Crushing  and  Roasting. — The  coarsely  crushed  ore  from  any  of  the 
storage-bins  is  drawn  off  as  needed  to  the  feed-hopper  of  the  dryer.  It  is 
dried  and  fine-crushed. 

Cripple  Creek  ore  is  roasted  in  a  mechanical  furnace,  such  as  the 
Edwards,  Figs.  74  and  75.  The  finishing  temperature  should  not  be  higher 
than  necessary  to  break  up  the  sulphates  formed  in  roasting. 

The  cooled  ore  is  raised  by  an  elevator  to  storage-bins,  whence  it  is 
drawn  as  needed  to  the  chlorination  barrels. 

The  Chlorination  Barrel. — This  is  shown  hi  perspective  in  Fig.  93,  and 
in  transverse  and  longitudinal  section  in  Fig.  94. 

Within  the  barrel  for  a  filter  a  perforated  2-in.  plank  floor  is  used. 
On  the  perforated  floor  rests  a  lead  sheet  of  4  Ib.  per  sq.  ft.  with  0.05-in. 
-holes,  f  in.  between  centers.  To  hold  down  the  filter-sheet,  a  wooden 
frame  or  grating  is  placed  upon  it.  This  is  held  by  blocks  h,  and  heavy 


138 


CHLORINATION  OF  GOLD  ORES 


strips  i,  securely  bolted  to  the  barrel.  The  wood  frames  beneath  last 
three  months;  those  above,  but  two  or  three  weeks.  This  wood-work, 
if  immersed  in  boiling  tar  or  asphalt  until  thoroughly  impregnated,  lasts 
longer  and  absorbs  but  little  solution.  The  common  size  of  barrel  is  6  ft. 
diameter  by  12  ft.  long,  and  the  capacity  is  8.5  tons. 

Charging  the  Roasted  Ore. — Into  the  cylinder  is  run  800  gal.  water, 
the  charge  of  8  tons  and  8  Ib.  of  liquid  chlorine.  Chlorine  can  be  obtained 
in  this  form  in  strong  steel  cylinders  or  drums.  The  charge-openings  of 


FIG.  93. — Chlorination  Barrel. 

the  barrel  are  now  closed,  and  it  is  rotated  at  the  rate  of  12  R.P.M.  for  a 
period  of  three  hours.     The  chlorine  roasts  thus : 


(1) 


Au+3Cl+H2O 


H2O. 


To  see  if  the  saturation  with  chlorine  is  complete  the  stopcojpk  j  is 
opened  and  the  issuing  gas  is  tested  with  ammonia,  which  produces  a  white 
fume  with  chlorine. 

If  needed  the  barrel  is  stopped  and  more  chlorine  is  added. 

The  precipitate  collects  upon  the  bottom  of  the  tank,  and  after  several 
charges  have  been  treated,  the  united  precipitate  is  drawn  off  at  D  and 
delivered  through  the  man-hole  L  into  the  pressure  tank  z  by  the  hose  y. 
The  precipitating  tank  is  then  washed  clean  with  the  aid  of  a  hose.  The 
pressure-tank  is  4  ft.  diameter  by  4i  ft.  high.  When  charged  the  cover  L 
is  clamped  in  place,  and  compressed  air,  under  a  pressure  of  40  Ib.  per  sq.  in., 


CHLORINATION  BARREL 


139 


~ 


140  CHLORINATION  OF  GOLD  ORES 

is  admitted  through  t.  At  the  same  time  connection  is  made  to  the  filtei 
press  T  through  the  pipe  u,  and  the  precipitate  collects  in  the  press  unde 
the  above  pressure.  This  filter-press  is  more  clearly  illustrated  in  Fig 
112.  The  filtrate  from  the  press  passes  over  a  sawdust  filter-bed,  as  i 
safeguard  before  it  is  run  to  waste.  The  sawdust  is  collected  occasionally, 
and  burned,  to  recover  the  small  amount  of  gold  which  it  may  have  caught 

The  precipitate  of  gold  sulphide  also  contains  sulphur,  and  sulphides  of 
arsenic,  antimony,  copper,  and  silver,  forming  a  "  sulphide  cake."  The 
press  is  next  opened  and  the  precipitate  withdrawn. 

Compressed  air,  entering  by  the  pipe  w  and  the  valve  c,  drives  the  gas 
through  the  pipe  v  and  the  lead  pipe  r  into  the  solution  in  the  tank,  and 
precipitates  the  gold  as  follows : 

(2)  2AuCl3+3H2S  =  Au2S3+6HCL 

The  gold  is  thus  thrown  down  as  an  auric  sulphide  in  a  solution  con- 
taining both  sulphuric  and  hydrochloric  acids.  The  reaction  is  rapid, 
taking  about  ten  minutes. 

At  first,  £[28  is  oxidized  by  the  chlorine,  thus: 

(3)  H2S+8C1+4H2O  =  H2S04+8HC1. 

Sulphuric  and  hydrochloric  acids  are  formed  by  the  reaction,  after  which 
auric  sulphide  is  precipitated  as  in  Equation  (2) . 

Filtering.— After  being  precipitated,  Au2Sa  is  allowed  to  settle  two 
hours.  The  clear  solution  then  is  drawn  off  at  C,  10  in.  above  the  bottom 
of  the  tank,  through  the  pipe  n  into  the  filter-press.  This  is  done  to  recover 
any  possible  flakes  of  gold  sulphide  that  failed  to  settle  in  the  tank  x. 
In  three  or  four  hours  after  precipitation,  the  tank  can  receive  a  fresh 
charge  of  gold-bearing  solution. 

After  saturation  with  chlorine,  the  barrel  is  revolved  for  an  hour, 
then  stopped  in  position  for  filtering  with  the  filtering  floor  down  and 
level.  The  outlet  pipe  k  is  connected  by  a  hose  to  the  settling  tank  and 
opened ;  and  water  is  pumped  into  the  barrel  above  the  charge  through  the 
valve  j.  The  solution  is  now  drained  off,  and  the  water  above  the^  charge 
forced  rapidly  through  by  means  of  compressed  air  introduced  through  the 
valve  j.  The  excess  of  chlorine  is  absorbed  by  the  wash-water,  and  does 
not  enter  the  building.  The  operation  of  filtering  is  next  suspended,  con- 
nections are  broken,  valves  closed  and  the  barrel  revolved  a  few  times  to 
mix  the  contents  again,  and  to  break  up  channels  that  may  have  formed 
during  the  leaching.  The  barrel  is  then  stopped,  water  run  in,  compressed 
air  admitted,  and  the  washing  resumed.  This  is  repeated  until  no  gold  is 
found  in  the  escaping  filtrate  when  tested.  The  compressed  air  admitted 
is  under  a  pressure  of  40  Ib.  per  sq.  in.  The  time  of  filtering  and 


CHLol!  I  NATION    PLANT 


141 


washing  on  an  average  is  2£  hours.  The  water  used  is  50  per  cent  the 
weight  of  the  ore.  All  connections  are  finally  broken,  valves  closed,  and 
manholes  opened,  and  the  cylinder  is  revolved  several  times  to  discharge  the 
contents.  It  is  then  washed  out  with  a  hose  to  prepare  it  for  another 
charge. 

Concentration. — The  washed  tailings  are,  before  being  discarded,  sub- 


FIG.  95. — Precipitation  Plant  for  Barrel  Chlorination. 


jected  to  a  table  concentration  to  recover  any  particles  of  unroasted  sul- 
phides or  tellurides  which  would  contain  gold. 

Clarifying. — The  filtrate  is  run  through  a  clarifying  press  and  sent  to 
the  stock  tanks.  From  this  it  is  pumped  through  the  opening  A ,  Fig.  95, 
into  the  precipitation  tank,  X,  which  is  10  ft.  diameter  by  12  ft.  high. 

Precipitation. — We  are  now  ready  to  precipitate  the  gold  by  passing  in 
H_-S.  To  do  this,  the  pipe  v  is  connected  to  the  lead-lined  generator  G 
which  contains  lumps  of  iron  sulphide  resting  on  a  perforated  lead  false- 
bottom.  Dilute  sulphuric  acid  admitted  below  the  false-bottom  comes  in 


142  CHLORINATION  OF  GOLD  ORES 

contact  with  the  iron  sulphide  and  abundantly  generates  H^S  according  to 
the  equation : 

(4)  FeS+H2SO4  =  FeS04+H2S. 

This  method  of  treatment,  successfully  conducted  at  the  Colorado 
Springs  plant,  was  abandoned  in  favor  of  the  cyanide  method  in  1912. 

When  the  ore  is  refractory  and  requires  preliminary  roasting,  this  cost 
adds  much  to  that  of  treatment.  In  chlorination,  roasting  is  common,  and 
in  any  case  it  improves  the  condition  of  the  ore,  and  makes  it  porous  and 
so  permeable  when  leached  or  filter-pressed. 


CHAPTER  XV 
(2)     CYANIDING    OF    GOLD    ORES 

This  is  a  hydrometallurgical  method  of  treatment,  that  is,  the  gold 
is  recovered  in  a  weak  water  solution  of  cyanide. 

GRAVITY  CONCENTRATION  PRIOR  TO  CYANIDING 

This  precedes  cyanide  treatment,  on  account  of  being  more  conveniently 
applied  at  this  point,  but  in  many  cases  it  would  be  advantageous  if  it 
were  feasible  to  have  it  follow  cyanide  treatment.  When  it  precedes 
cyanide  treatment,  a  considerable  proportion  of  the  readily  soluble  gold 
and  silver  may  be  removed  in  the  concentrate,  when  it  might  be  more 
advantageous  to  recover  them  as  bullion.  In  the  case  of  low  freight  rates 
and  favorable  smelter  contracts  this  may  be  an  advantage  rather  than 
otherwise ;  but  for  the  majority  of  plants  the  more  gold  and  silver  turned  out 
as  bullion,  the  better.  Therefore,  the  ideal  practice  in  most  cases  would  be 
to  recover  all  the  gold  and  silver  possible  as  bullion  by  the  cyanide  process 
and  then  concentrate  the  tailing  to  recover  as  high  a  percentage  of  the 
remaining  gold  and  silver  which  had  escaped  dissolution  as  feasible.  Obvi- 
ously there  are  many  cases  where  the  percentage  of  extraction  of  gold  and 
silver  from  the  residue  leaves  no  margin  for  concentration. 

At  times  concentration  can  be  introduced  into  the  cyanide  flow  sheet 
to  advantage  for  the  purpose  of  removing  the  minerals  which  are  difficult, 
if  not  impossible,  to  treat  by  the  regular  milling  scheme,  so  that  they  may 
receive  the  special  treatment  necessary  without  having  to  incur  the  expense 
of  subjecting  the  whole  tonnage  to  the  special  treatment  made  necessary 
by  a  comparatively  small  proportion  of  refractory  minerals.  On  the  other 
hand,  the  introduction  of  concentration  adds  to  the  first  cost  of  the  plant 
and  the  complexity  of  its  operation,  so  that  these  disadvantages  must  be 
carefully  weighed  against  the  advantages  which  are  likely  to  accrue. 

OUTLINE  OF  THE  PROCESS  OF  CYANIDING 

The  ore  is  crushed  to  such  fineness  that  its  contained  gold  is  left  open 
to  the  action  of  a  weak  solution  of  potassium  or  sodium-cyanide.  The 
gold,  brought  into  solution,  is  then  precipitated  from  the  clear  filtrate  and 
this  precipitate  melted  into  form  of  a  bar  or  ingot.  The  gold  in  the 

143 


144  CYANIDING  OF  GOLD  ORES 

cyanide  solution  is  present  as  potassium  or  sodium  auro-cyanide.  With 
certain  refractory  ores  the  solvent  power  of  the  solution  is  greatly 
increased  by  the  addition  of  bromine  or  bromo-cyanogen. 

Extraction  of  gold  by  cyaniding  is  successful,  where  gold  cannot  be 
completely  removed  by  amalgamation.  This  is  the  case  with  pyrite  ores, 
and  such  extraction  can  be  practiced  to  advantage,  not  only  where  amalga- 
mation is  unsuitable,  but  where  smelting  is  too  expensive. 

Practically  all  gold  or  silver  ores  can  be  treated  by  the  cyanide  process 
or  by  the  cyanide  process  in  combination  with  some  other  preliminary  or 
accessory  treatment,  y/ith  the  exception  perhaps  of  certain  ores  containing 
a  considerable  amount  of  copper,  lead,  etc.  In  such  cases,  the  desirability 
of  recovering  the  base  metals  leads  to  smelting  the  ore  upon  either  the  lead 
or  copper  basis,  and  the  necessity  for  hydrometallurgical  treatment  for  the 
recovery  of  the  gold  and  silver  disappears  since  the  precious  metals  are 
recovered  by  the  smelting  operation  in  connection  with  the  base  metals. 

Cyanidation  has  proved  a  remarkably  cheap  and  efficient  method  of 
extraction,  but  it  has  limitations,  not  only  in  respect  to  the  solubility  of 
the  gold,  but  because  of  the  action  of  certain  compounds,  notably  those 
of  copper,  that  are  sometimes  present  and  interfere  with  extraction  in 
various  ways.  The  .process  has  the  advantage  over  the  chlorination 
method  in  that  silver,  as  well  as  gold,  can  be  extracted.  Under  favorable 
conditions  the  extraction  is  high,  and  modern  methods  have  reduced 
the  cost  of  treatment  to  a  low  figure. 

The  Use  of  Hot  Solutions  in  Cyaniding  for  Better  Extraction  of  Silver 
and  Gold. — At  the  Belmoiit  mill,  Tonopah,  the  temperature  at  the  stamps 
is  60°  to  70°  F.,  and  by  the  use  of  exhaust  steam  at  the  Pachuca  agitators, 
the  pulp  is  raised  to  90°  to  100°  F.,  resulting,  as  reported,  in  increased 
extraction  over  using  cold  solutions  of  2  per  cent.  At  the  Montana-Tono- 
pah  mill  crushing  is  done  at  50°  to  60°  F.,  and  by  live  steam  in  the  agitator, 
the  temperature  is  brought  up  to  1 10°  F.  It  is  found  at  this  mill  that  when 
heating  is  not  done  extraction  falls  off,  also  that  heat  aids  settling.  At  the 
MacNamara  mill,  Tonopah,  the  pulp  was  heated  to  115°  to  120°  F.,  using 
live  steam  in  the  agitators,  whereby,  as  compared  with  cold  solutions  the 
extraction  increased  and  the  time  of  agitation  was  lessened.  E^en  as 
compared  with  a  temperature  of  80°  an  increase  to  120°  improved  extrac- 
tion by  1.5  per  cent  to  2  per  cent.  The  cost  of  this  heating  may  be  reckoned 
at  18  to  30  cents  per  ton  as  against  a  saving  at  2  per  cent  or  60  cents  per 
ton  on  dollar  silver.  On  gold  ore  savings  are  not  so  secured.  At  Kal- 
goorlie,  Western  Australia,  where  ores  are  roasted,  the  temperature,  after 
mixing  with  solution,  is  as  high  as  200°  F.,  but  they  prefer  to  cool  the  ore 
before  mixing. 


CYANIDING  ORES  145 


ORES  SUITED  TO  CYANIDATION 

Dealing  with  gold-bearing  ores,  the  following  classes  are  amenable  to 
treatment : 

1.  Talcose  or  Clayey  Ores. — When  crushed, 'these  produce  a  high  per- 
centage of  slime,  which  is  too  fine  for  leaching,  and  gives  trouble  in  any 
type  of  filter.     Generally,  the  capacity  of  a  plant  treating  these  ores  is' 
not  high. 

2.  Free-milling   Silicious   Ores. — These   constitute  the   bulk   of  the 
ores   treated   by  the   cyanide    process  throughout  the  world.     The  gold 
may  be  fine  or  coarse,  in  the  latter  case  being  removed  by  amalgamation 
before  subsequent  cyanide  treatment.     It  is  not  economical  to  dissolve 
coarse  gold  in  cyanide. 

3.  Pyrite  Ores. — Gold  in  this  class  is  in  most  cases  mechanically  mixed 
with  the  iron  pyrite,  which  when  crushed,  liberates  the  gold  for  the  solu- 
tion in  cyanide.     It  has  been  found  in  several  mining  districts  that  it  is 
possible  to  treat  the  pyrite  mixed  with  the  ore,  no  concentration  being 
necessary;  but  in  the  majority  of  cases,  it  is  found  better  to  concentrate, 
and  treat  this  product  separately. 

4.  Telluride  Ores. — These  occur  at  Cripple  Creek,   Goldfield,   Kal- 
goorlie,  and  other  places,  and  while  not  complex,  have  given  considerable 
trouble  in  treatment.     Ordinary  cyanide  solutions  are  not  effective  on 
tellurium  compounds,  and  the  ore  should  be  either  concentrated  and  treated 
by  ordinary  cyanide  or  bromo-cyanide,  or  all  the  ore  roasted,  followed  by 
the  usual  cyanide  methods. 

5.  Antimony  Ores. — Antimony  is  a  troublesome  mineral  to  deal  with 
in  gold  extraction.     It  occurs  in  gold  ores  in  Rhodesia  to  some  extent,  and 
in  New  South  Wales.     Roasting  seems  to  be  effective,  while  caustic  soda 
solutions  have    helped    cyanidation.     Tailing  containing  antimony  has 
been  successfully  treated  in  Australia  with  no  special  process. 

6.  Graphite  Ores. — At  Ashanti,  West  Africa,  Kalgoorlie  and  Gympie, 
Australia,  graphitic  slate  or  schist  is  mixed  with  the  ore,  and  while  not 
containing  a  high  percentage  of  gold  it  causes  a  premature  precipitation  of 
gold  from  cyanide  solutions. 

7.  Copper  Ores. — Many  gold  ores  contain  a  low  percentage  of  copper, 
and  with  care  may  be  treated  with  a  fair  recovery.     Copper  gradually 
changes  cyanide  solutions,  and  then  is  precipitated  on  the  zinc  shaving, 
preventing  a  proper  precipitation  of  gold.     Copper  in  the  resultant  bullion 
with  care  may  be  refined. 

8.  Arsenical  Ores. — A  mineralized  ore  often  contains  a  little  arsenic, 
but  pure  mispickel,  or  arsenical  pyrite,  requires  skill  in  treatment.     The 
ore  may  be  roasted  after  fine  crushing,  or  concentrated,  and  the  product 
roasted  alone.     Noted  cases  of  mines  producing  this  ore  are  at  Bendigo, 


146  CYANIDING  OF  GOLD  ORES 

in  Victoria;  the  Lancefield  and  Transvaal  mines,  in  Western  Australia; 
the  Deloro  and  Hedley  in  Canada.  From  the  first  mentioned  group,  the 
pyrite  is  roasted  and  treated  by  cyanide  or  chlorination ;  at  the  Lancefield 
both  wet  and  dry  processes  have  been  tried  on  large  tonnages;  the  Trans- 
vaal mine  ore  is  very  refractory;  while  at  the  Deloro,  bromo-cyanide  was 
used  for  several  years.  At  the  best,  it  may  be  said  that  such  a  gold-bearing 
ore  is  difficult  to  treat. 

CHEMISTRY  OF  THE  CYANIDE  PROCESS  FOR  GOLD  ORES 

When  a  solution  containing  from  0.1  to  0.5  per  cent  cyanide  is  brought 
into  contact  with  crushed  ore  containing  very  fine  gold,  this  metal  is  easily 
dissolved.  According  to  Eisner,  the  equation  is  as  follows: 

(1)  2Au+4KCN+O+H2O  =  2AuK(CN)2+2KOH. 

The  gold  is  dissolved  by  the  action  of  potassium  or  sodium  cyanide  in  the 
presence  of  oxygen  and  water,  forming  an  anric-potassic  cyanide  and  caustic 
potash.  Oxygen  is  needed  to  fulfill  the  requirements  of  the  reaction,  and 
consequently  ore,  or  solution  acting  on  ore,  must  be  aerated  in  some  man- 
ner. When  oxygen  of  the  dissolved  air  is  consumed,  action  ceases,  but 
resumes  with  a  fresh  supply  of  air.  Oxidizing  agents  such  as  potassium 
chlorate  and  permanganate,  and  the  peroxides  of  lead,  manganese,  sodium, 
and  barium  may  be  used  to  furnish  oxygen  in  place  of  air,  but  have  been 
found  too  expensive  for  practical  use. 

The  inadequacy  of  this  equation  as  a  guide  to  consumption  of  cyanide 
is  at  once  apparent  in  the  treatment  of  complex  ores,  with  their  various 
constituents.  The  equation  calls  for  1.51  units  of  Au  or  0.83  unit  of  Ag 
per  unit  of  KCN.  With  many  silver  ores  this  relation  is  closely  approxi- 
mated, but  with  gold  ores  a  consumption  of  20  of  KCN  to  1  of  Au  is  con- 
sidered satisfactory,  and  40  to  1  is  more  common.  The  mixed  potassium 
cyanide  salt  (98  to  99  per  cent  KCN)  has  been  quite  generally  sup- 
planted by  sodium  cyanide  (120  to  129  per  cent,  in  terms  of  KCN). 
Subsequent  references  to  cyanide  will  be  in  terms  of  100  per  cent  KCN, 
although  sodium  cyanide  is  used.  t 

When  an  ore  containing  pyrite  is  exposed  to  the  weather,  air  and  mois- 
ture slowly  act  on  the  mineral,  with  the  following  reaction : 

(2)  3FeS2+2H20+ 22O  =  FeSO4+Fe2(SO4)3+2H2SO4. 

Ferrous  and  ferric  sulphates  and  sulphuric  acid  are  thus  formed.  The 
first  two  named  would  tend  to  precipitate  gold.  Ferric  sulphate  is  acid  in 
its  reaction,  and  with  sulphuric  acid,  if  not  neutralized,  it  would  decompose 
and  cause  a  serious  loss  of  cyanide.  Such  compounds  are  called  "  cyani- 
cides."  Ore,  therefore,  which  contains  pyrite,  and  has  been  exposed  to  the 


.REACTIONS  IN  CYANIDING  147 

weather  needs  caustic  soda  or  lime  to  neutralize  the  acidity,  and  an  excess 
to  provide  for  any  acidity  resulting  from  further  decomposition.  This 
excess  is  termed  the  "  protective  alkalinity."  To  remove  the  soluble  ferrous 
sulphate  and  sulphuric  acid,  water-wash  before  treatment  should  be  suffi- 
cient, but  would  require  some  time,  so  a  certain  quantity  of  lime  is  added. 
Lime  is  cheaper  than  and  preferable  to  caustic  soda,  the  latter  making 
undesirable  compounds,  often  noticed  later  in  treatment.  In  some  dis- 
tricts, the  question  of  freight  will  decide  which  should  be  used.  The  action 
of  lime  is  as  follows : 

(3)  FeS04+Ca(OH)2  =  Fe(OH)2+CaSO4 

(4)  Fe2(S04)3+3Ca(OH)2  =  2Fe(OH)2+3CaS04 

(5)  H2SO4+Ca(OH)2  =  2H2O+CaSO4 

The  result  is  the  formation  of  a  harmless  iron  hydroxide  arid  calcium  sul- 
phate. 

The  reaction  that  takes  place,  when  a  gold-bearing  solution  comes  hi 
contact  with  zinc-shaving  in  the  precipitating  boxes,  or  when  zinc-dust  is 
mixed  with  it  is: 

(6)  KAu(CN)2+2K(CN)2+ Zn-f  H2O  =  K2Zn(CN)4+Au+KOH+H 

in  which  one  part  of  zinc  is  computed  to  precipitate  three  parts  of  gold. 
The  gold  forms  a  brown  or  black  precipitate  and  the  zinc  potassic  cyanide 
remains  in  solution. 

Cyanide  is  also  used  up  by  direct  combination  with  the  zinc  as  follows: 

(7)  Zn+4K(CN)  +2HO2O  =  K2Zn(CN)4+2KOH  +H2. 

In  both  the  foregoing  reactions  hydrogen  escapes  in  bubbles. 
When  aluminum  is  used  we  have : 

(8)  KAu(CN)2+2K(CN)2+2Al+H2O 

=  K2Al(CN)+42Au+K2Al2O4+4H, 

in  which  one  part  of  aluminum  precipitates  7.2  parts  of  gold. 

The  barren  solution,  from  which  the  gold  has  been  extracted,  is  used 
again  in  the  mill,  and  accumulates  impurities  from  ore  that  is  being  treated, 
and  from  the  zinc  with  which  it  was  in  contact  in  the  zinc-boxes.  As  a 
result,  it  gradually  becomes  less  efficient  than  fresh  solution.  It  has  been 
found  that,  on  adding  lime  to  a  cyanide  solution,  its  solvent  power  upon  a 
clean  ore  is  increased,  but  not  on  a  sulphide  ore.  Such  a  solution,  if  treated 
with  sodium  sulphide  to  the  point  of  exact  neutrality,  and  with  a  small 


148  CYANIDING  OF  GOLD  ORES 

excess  of  lead  acetate,  and  given  time  to  permit  the  resultant  sulphide  to 
precipitate,  is  improved  in  dissolving  power  as  follows: 

-(9)  K2Zn(CN)4+Na2S  =  K2Na2(CN)4+ZnS. 

The  cyanide  is  here  regenerated,  while  the  zinc  sulphide  separates.  This 
is  a  means  of  overcoming  the  accumulation  of  zinc  in  solution,  which  is  one 
of  the  drawbacks  to  the  use  of  zinc  for  precipitation,  compared  with  elec- 
trical deposition.  Chemicals,  however,  are  not  indispensable  for  disposing 
of  the  zinc.  In  Eisner's  equation,  caustic  potash  is  set  free.  This  reacts 
upon  the  sulphides  in  an  ore,  forming  soluble  sulphides,  which  in  turn  react 
like  the  sodium  sulphide  in  the  reaction  above,  precipitating  zinc  sulphide 
from  the  ore. 

The  following  minerals  and  chemical  compounds  destroy  or  combine 
with  cyanide,  and  render  it  incapable  of  dissolving  gold:  Copper  in  the 
form  of  .sulphate,  carbonate,  copper  glance,  erubescite,  or  copper  pyrite. 
(The  sulph-antimonites  of  copper  are  without  action.)  Manganese  as 
"  wad  "  (impure  hydrous  oxide),  but  not  the  carbonate  or  oxide;  zinc 
as  smithsonite,  but  not  blende  or  zinc  silicate. 

Graphite,  which  is  found  in  certain  ores,  also  carbon  remaining  in  burned 
lime,  both  interfere  with  extraction  by  causing  a  premature  precipitation  of 
gold.  Also  leaves,  roots,  and  other  organic  matter  act  in  the  same  way. 

It  has  been  found  that  gold  thus  prematurely  precipitated  is  soluble  in  a 
solution  of  sodium-sulphide,  this  being  added  when  leaching  the  sands  after- 
practically  all  the  cyanide  has  been  displaced  by  a  water-wash.  Precip- 
itation of  the  gold  is  effected  by  passing  the  Na2S  through  boxes  filled  with 
copper  shaving,  but  little  copper  going  into  solution. 

To  increase  the  activity  of  zinc  shaving  in  precipitating  gold,  it  may  be 
dipped  in  10  per  cent  lead  acetate  solution,  or  a  drip  of  the  latter  may 
be  fed  in  at  the  head  of  the  zinc-boxes.  This  forms  a  zinc-lead  couple, 
which  reacts  electrically  on  the  gold  solution.  Both  potassium  and 
sodium  cyanide,  98  and  128  per  cent  pure,  are  used  with  varied  results, 
the  latter  being  rather  more  favored  as  a  dissolving  agent. 

The  cyanogen  contents  of  potassium  and  sodium  cyanide  respectively 
are  about  38  and  51  per  cent,  so  for  the  same  weight  of  salt  there  is  theiextra 
percentage  of  cyanogen,  a  consideration  where  freight  is  costly.  Com- 
mercial cyanide  contains  small  quantities  of  alkaline  sulphides,  whose 
presence  diminishes  the  solvent  power  of  the  cyanide.  It  pays,  therefore, 
to  buy  the  salt  on  a  guaranteed  analysis. 

The  Concentration  of  the  Cyanide  Solution  is  a  vital  point.  In  general, 
the  stronger  the  solution,  within  certain  fairly  well-defined  limits,  the  more 
rapid  the  dissolution,  and,  furthermore,  the  less  the  interference  of  a  given 
percentage  of  impurity  which  might  be  present  in  solution.  On  the  other 
hand,  a  greater  proportion  of  impurity  may  be  dissolved  by  the  stronger 


THE  BROMO-CYANIDE  PROCESS  149 

solution.  In  the  case  of  a  plant  which  is  operating  at  forced  capacity  a 
stronger  solution  may  be  used  to  advantage  in  order  that  the  maximum 
extraction  may  be  attained  under  this  condition  of  operation.  Stronger 
solutions  may  result  in  increased  cyanide  consumption,  although  the  mag- 
nitude of  this  loss  in  a  properly  operated  plant  is  not  so  great  as  has  been 
supposed. 

However,  in  a  plant  where  there  is  considerable  mechanical  loss  of  solu- 
tion the  additional  cyanide  loss  would  prove  an  important  factor.  For 
this  reason  a  strong  solution  is  not  generally  favored  when  continuous 
decantation  is  used.  The  tendency  in  some  cyanide  plants  has  been  to  use 
a  lower  concentration  in  cyanide  than  that  capable  of  giving  the  highest 
economical  result.  This  probably  is  through  the  fear  of  excessive  cyanide 
loss. 

The  Bromo-cyanide  Process.  —  Tellurides  of  gold  and  silver  are  prac- 
tically insoluble  in  plain  cyanide,  but  are  soluble  in  bromo-cyanide  solu- 
tions. The  latter  process  was  first  used  on  a  large  scale  in  1899  at  the 
Hannans  Star  mill,  Kalgoorlie,  which  was  mainly  a  customs  plant  receiving 
many  shipments  of  rich  telluride  ores.  The  process  involved  is  commonly 
known  as  the  Diehl. 

The  bromo-cyanide  solution  is  made  according  to  the  following  equation  : 


(10) 

Its  action  in  the  treatment  vat  is  supposed  to  be  as  follows  : 
(11)  BrCN+3KCN+2Au  =  2KAuCN2+KBr. 

The  first  two  quantities  in  Equation  (10)  are  contained  in  the  mixed 
salts  having  about  40  to  44  per  cent  Br  as  KBr,  and  20  to  22  per  cent  Br 
as  KBrOa;  the  proportion  of  Br  as  bromide  being  about  twice  that  of  Br 
as  bromate.  A  30-lb.  charge  is  usually  made  up,  50  Ib.  of  63  per  cent 
H2SO4,  20  Ib.  of  KCN  of  93  per  cent  and  36.8  Ib.  of  mixed  salts  as  above 
given. 

The  solution  is  made  in  a  closed  wooden  vessel,  holding  about  200  gal., 
stirred  by  rotating  arms.  In  making  up  a  charge,  a  portion  of  the  water 
and  all  the  H2SO4  are  first  mixed,  and  allowed  to  cool  to  normal  tempera- 
ture. The  KCN,  which  is  dissolved  in  a  separate  vessel  in  sufficient  water 
to  fill  the  mixing  vessel,  is  then  run  in,  and  at  the  same  time  the  proper 
weight  of  "  mixed  salts  "  is  gradually  added.  The  whole  is  then  agitated 
for  six  hours  before  being  used,  and  in  a  closed  vessel  it  will  retain  its 
strength  for  some  days.  The  cost  of  a  30-lb.  charge  of  BrCN  is  about 
$21.60. 

Experiments  have  shown  the  following  points  necessary  for  good  work: 

1.  The  daily  ore  sample  should  be  taken  in  the  morning,  and  assayed 


150  CYANIDING  OF  GOLD  ORES 

as  soon  as  possible,  so  that  the  value  of  the  ore  passing  to  the  vats  in  the 
previous  twenty-four  hours  may  be  determined. 

2.  The  pulp  should  have  a  long  KCN  treatment. 

3.  A  vat  should  be  kept  under  KCN  treatment  until  the  value  of  the 
KCN  residue  is  known. 

4.  The  alkalinity  of  the  vat  should  then  be  determined  and  corrected 
to  0.01  per  cent  by  H2SO4  before  adding  BrCN. 

5.  The  quantity  of  BrCN  added  should  then  be  determined  from  the 
value  of  KCN  residue,  and  the  tonnage  of  the  vat. 

6.  The  lime  added  to  the  ore  during  crushing  should  be  varied  accord- 
ing to  the  alkalinity-test  after  KCN  treatment,  so  that  the  plant-solution 
tests  about  0.02  per  cent. 

7.  Lime  water  should  be  made  and  added  to  the  vats  or  to  the  solution 
from  the  presses,  instead  of  adding  lime  to  the  vats. 

8.  Metallic  iron  should  be  kept  out  of  the  pulp  as  far  as  possible, 
as  it  is  both  a  cyanicide  and  a  bromo-cyanicide. 

In  operation,  a  vat,  when  full,  was  given  its  charge  of  KCN,  and  three 
hours  afterward  a  "  dip  "  KCN  residue  was  taken,  and  the  charge  of 
BrCN  solution  added.  After  a  total  agitation  of  twenty  hours  a  quantity 
of  lime  was  added,  and  the  vat-charge  "  pressed."  The  quantity  of  BrCN 
added  was  varied  according  to  the  residue  of  preceding  vats,  and  the  value 
of  the  ore  being  tireated  as  shown  by  the  daily  ore-sample.  Each  charge  of 
bromo-cyanide  was  totally  destroyed.  In  places  where  fuel  and  furnace 
supplies  were  expensive,  this  chemical  process  would  show  a  decided  advan- 
tage. The  process  requires  more  metallurgical  skill  and  constant  attention 
to  the  progress  of  each  vat  under  treatment;  but,  if  this  is  available,  the 
results  are  highly  satisfactory,  and  the  process  has  definite  claim  to  be  a 
cheap  and  efficient  method  of  treatment  for  such  ores  as  the  Kalgoorlie 
sulpho-tellurides. 

THE  STANDARD  SYSTEMS  OF  CYANIDATION 

These  may  be  divided  into  three  as  follows : 

1.  Sand  leaching,  where  the  whole  ore  pulp  after  grinding  is  classified 
into  two  products,  sand  and  slime,  the  sand  being  sent  to  percolation^tanks 
for  leaching;  the  slime  or  fine  product  being  filtered  or  decanted  to  yield 
a  pregnant  solution  and  a  solid  residue. 

2.  Filter  slime  treatment,  where  the  whole  pulp  is  ground  fine  and  fil- 
tered to  yield  a  pregnant  solution  and  a  solid  residue. 

3.  Slime  Agitation,  where  the  whole  pulp  is  ground  fine  and  agitated 
for  a  considerable  period,  then  by  countercurrent  decantation  yields  a 
pregnant  solution  and  a  barren  or  nearly  barren  residue  that  is  thrown 
away. 

Practically  the  "  sand  "  is  the  portion  of  the  pulp  treated  by  percola- 


SAND  LEACHING  151 

tion  or  leaching ;  while  the  balance  is  considered  to  be  slime  and  is  treated 
by  suction  or  pressure  filtration. 

With  the  important  exceptions  of  the  Rand  and  the  Homestake  the 
earlier  practice  of  sand  leaching,  that  is  of  comparatively  coarse  grinding 
and  treating  the  sand  and  slime  separately  has  largely  given  way  to  the 
present  practice  of  (2)  filter  slime  treatment  or  (3)  slime  agitation  (both 
"  all-sliming  "  treatment)  the  pulp  being  finely  ground  and  treated  as  a 
single  product. 

Referring  back  to  the  so-called  sand  leaching,  the  pulp,  rather  coarsely 
ground,  is  classified  hi  cone  classifiers  such  as  the  Caldecott  giving  a  sand 
for  percolation  and  a  slime  such  that  90  per  cent  of  it  will  pass  200-mesh 
(0.0029  in.). 

In  all-sliming  plants  the  pulp  varies  from  60  to  90  per  cent  through 
200-mesh,  depending  on  the  economical  limit  of  fine  grinding,  and  subse- 
quent treatment.  In  sand  and  slime  plants,  crushing  is  almost  universally 
done  in  water,  while  with  all-sliming,  crushing  in  cyanide  solution  is  almost 
universal. 

(1)      SAND  LEACHING 

Separation  of  Sand  from  Slime. — This  is  an  interesting  problem,  and 
is  performed  either  to  furnish  from  the  pulp  a  sand  for  fine  grinding  or 
again  to  make  two  products,  sand  and  slime,  the  sand  to.be  leached  into 
vats,  the  slime  to  be  separately  treated.  The  machines  used  in  separating 
the  sand  from  the  slime  and  their  principles  of  operation  are  fully  described 
under  head  of  "  Classifiers." 

Leaching  the  Sands. — Cyanidation  was  first  applied  to  the  recovery 
of  gold  in  accumulated  tailings  from  stamp-mills.  In  South  Africa  this 
material  still  containing  $3.50  per  ton,  was  impounded  or  retained  behind 
dams  until  cyaniding  could  be  undertaken.  The  tailing  was  shoveled  into 
cars  and  hauled  to  large  leaching  vats.  Here  the  ore  was  leached  with 
weak  cyanide  solution,  the  gold  precipitated  from  the  filtered  solution  by 
passing'  it  through  boxes  containing  zinc  shavings.  The  precipitate  was 
treated,  melted,  and  obtained  in  form  of  gold  ingots  or  bars.  This 
practice,  as  long  as  the  impounded  tailings  lasted,  was  quite  simple,  but 
it  has  been  abandoned  with  the  exhaustion  of  those  accumulations. 

Description  of  Vats. — Leaching  vats  are  constructed  with  filter  bottoms 
and  may  be  made  of  wood  or  steel.  In  warm  countries,  like  Australia, 
South  Africa  and  Mexico  the  steel  vat  is  preferred ;  but  in  cold  countries, 
where  it  is  necessary  to  house  the  plant,  the  wooden  vat  gives  satisfaction. 
The  latter  is  cheaper  in  first  cost  and  easier  to  set  up,  though  the  wood 
absorbs  gold  solution.  The  steel  vat,  on  the  other  hand  is  less  liable  to 
leak,  but  should  be  painted.  A  wooden  vat,  also,  when  requiring  it, 
should  be  painted.  In  Western  America  wooden  vats  predominate. 


152 


CYANIDING  OF  GOLD  ORES 


Fig.  96  is  a  perspective  view  of  a  wooden  vat  with  the  filter-cloth 
omitted.  This  shows  the  false-bottom  of  slats  and  the  hinged  bottom- 
discharge  opening  through  which  the  exhausted  tailing  is  shoveled  or  sluiced 


FIG.  96.— View  of  Wooden  Leaching  Vat, 

out.  A  wooden  ring,  2  in.  high  and  2|  in.  thick,  is  nailed  to  the  bottom  of 
the  vat,  leaving  a  space  of  J  in.  between  it  and  the  side.  Parallel  strips, 
1  in.  high,  are  nailed  a  foot  apart  upon  the  bottom,  and  across  these  1-  by 


FIG.  97.— Plan  of  Steel  Leaching  Vat. 


4-in.  strips  or  slats  are  laid  with  1  in.  space  between.  Upon  this  false- 
bottom  cocoa-matting  is  spread,  and  over  it  8-oz.  canvas  filter  cloth  cut 
12  in.  larger  in  diameter  than  the  vat.  The  edges  of  the  cloth  are  held 


DOUBLE  TREATMENT 


153 


down  by  a  rope  laid  upon  the  canvas  and  driven  into  the  J-in.  space  between 
the  staves  and  the  wooden  ring. 

Figs.  97  and  98  represent  in  plan  and  in  elevation  respectively,  the  con- 
struction of  a  steel  vat  having  a 
perforated  board  bottom.  A  ring 
of  flat  iron  \  by  2J  in.  is  riveted 
to  the  side  of  the  vat,  with  space- 
thimbles  to  hold  it  \  in.  from  the 
side.  The  cleats  that  sustain  the 
false  bottom  are  2  in.  high  by 
\\  in.  wide.  The  1-in.  bottom- 
boards  are  bored  with  J-in.  holes 
and  screwed  to  the  cleats.  As 
in  the  case  of  the  wooden  vats, 
the  thick  stiff  cocoa-matting  is 
laid  upon  the  false-bottom,  and 


FIG.  98. — Section  of  Steel  Leaching  Tank. 


covered  with  a  filter-cloth  of  8-  to  10-oz.  canvas.  The  edges  are  calked  with 
J-in.  rope  into  the  J-in.  space,  as  shown  at  g  in  the  sectional  view,  Fig.  98. 

Leaching  vats  vary  in  size  from  16  to  50  ft.  in  diameter  and  4  to  9  ft. 
in  depth.  The  shallow  ones  are  for  the  more  finely  ground  sand. 

Double  Treatment. — Fig.  99  shows  the  two  steel  vats  or  intakes  used 
in  this  system  as  practiced  in  South  Africa.  It  consists  of  an  upper  or 
settling  vat  40  ft.  diameter  by  1\  ftl  high  to  which  the  sands  are  con- 
veyed to  be  spread  out  and  to  receive  a  preliminary  leaching,  and  of  a 
lower  or  leaching  vat  12  in.  deeper.  Both  vats  are  carried  on  a  steel 
structure  for  access  both  above  and  below.  When  the  leaching  is  suf- 
ciently  completed  in  the  settling  vat,  then  seven  bottom  doors  or  valves 
are  opened  and  the  material  is  shoveled  into  the  leaching  vat  beneath,  40 
ft.  by  8J  ft.  high.  When  thus  again  handled,  the  tailings  become  more 
bulky,  and  more  open  and  even  for  leaching.  Here  leaching  and  wash- 
ing is  completed.  The  exhausted  tailings  are  then  withdrawn  at  the 
discharge  openings  similar  to  those  of  the  upper  tank,  and  trammed 
away  to  the  waste  dump. 

Leaching  the  Sands. — While  transferring  the  sand  from  collectors  to 
leaching-vats,  lead  acetate,  previously  dissolved  in  water  (J  Ib.  per  ton)  is 
added  and  slacked  lime  (4  Ib.  per  ton)  is  thrown  into  the  collecting  vat,  thus 
thoroughly  mixing  the  lime  with  the  sand.  After  transferring,  a  little 
shoveling  is  done  to  level  the  sand.  The  first  leaching  solution,  amounting 
to  30  tons,  is  brought  up  to  0.25  per  cent  strength  by  the  addition  of  a 
sufficient  amount  of  potassium  cyanide  solution  of  known  strength  to  the 
vat  under  treatment.  This  amount  of  strong  solution  is  allowed  to 
drain  slowly  through  the  partly  opened  drain-valves  and  is  followed  by 
repeated  washings  of  weak  solution  from  0.15  to  0.20  per  cent,  after  which 


154 


CYANIDING  OF  GOLD  ORES 


the  charge  is  drained  for  transfer.     This  first  treatment  occupies  five  days, 
including  time  of  the  latter. 


-10  6 *j< 10  6- 

OUTSIDE    ELEVATION 


FIG.  99.— Vats  for  Double  Treatment. 
SIZING  TEST  ON  SAND  RESIDUE 


Screen  Mesh.          Percentage. 


Remaining  on 

20 

0  15 

Remaining  on  

30 

11  64 

Remaining  on  

40 

13  98 

Remaining  on.  .  .    . 

50 

12  31 

Remaining  on 

60 

10  48 

Remaining  on  

80 

17  54 

Remaining  on  

100 

12  77 

Passing  

100 

21  05 

• 

99.92 

SAND  LEACHING  155 

The  second  treatment  averages  five  days  and  consists  of  repeated 
washings  of  strong  and  weak  solutions,  that  are  drained  off,  and  the  sand 
transferred  to  another  vat  for  the  final  treatment,  which  consists  of  as 
many  washes  of  wash  solution  as  there  is  time  to  apply,  followed  by  two  or 
three  of  water  to  displace  all  the  solution.  Then  the  vat  is  finally  drained 
by  vacuum  for  discharging  from  the  plant.  Each  charge  of  solution  is 
allowed  to  disappear  below  the  surface  of  the  sand  before  the  succeeding 
one  is  applied.  Sand  undergoes  treatment  for  twelve  to  fifteen  days. 

All  teachings  from  the  sand  vats,  as  well  as  the  plant  solutions,  are 
sampled,  assayed,  and  titrated  for  cyanide  and  alkalinity  daily.  Attenu- 
ated leaching  solutions  are  sent  direct  to  weak  sumps.  Centrifugal  pumps, 
when  not  pumping  to  treatment-vats,  are  in  service  circulating  solution  in 
sumps  through  cones,  Fig.  58,  for  the  purpose  of  aerating. 

All  potassium  cyanide  used  in  the  treatment  of  sand,  is  dissolved  in  a 
small  vat  from  which  a  2-in.  pipe-line  is  connected  to  the  suction  of  a  4-in. 
centrifugal  pump  used  to  pump  solutions  on  sand.  By  means  of  a  table 
and  float  arranged  on  the  vat,  the  desired  strength  of  solution  can  be 
obtained  by  opening  the  2-in.  line  and  allowing  the  requisite  amount  of 
standard  solution  to  be  drawn  through  the  pump  with  the  weak  solution 
from  the  weak  sumps. 

The  Rand  is  the  greatest  present-day  exponent  of  sand  leaching, 
comparatively  few  plants  in  other  parts  of  the  world  retaining  separate 
sand  and  slime  treatment.  The  important  improvements  made  in  fine 
grinding  and  the  handling  of  slime  created  a  marked  tendency  toward  all- 
sliming  and  the  abandonment  of  sand  leaching,  this  trend,  in  a  few  instances 
resulting  in  inadequate  consideration  of  the  question  of  highest  commer- 
cial recovery  versus  theoretical  extraction.  A  few  all- sliming  plants,  after 
careful  research  and  consideration,  have  found  it  advisable  to  revert  to  the 
former  sand  and  slime  practice.  But  even  in  the  strongholds  of  sand  leach- 
ing, the  present  day  trend  is  toward  finer  grinding  and  a  gradual  reduction 
in  the  proportion  of  total  mill  pulp  treated  by  percolation. 

The  removal  of  the  residue  from  the  leaching  tanks,  which  now  range 
up  to  70  ft.  diameter  by  12  ft.  deep,  is  accomplished  by  sluicing,  shovel- 
ing, or  mechanical  excavators,  the  method  depending  on  local  factors  such 
as  cost  and  abundance  of  water,  and  labor  and  storage  requirements.  A 
recent  trend  in  the  method  of  sand-residue  disposal  has  been  to  sluice 
the  sand  into  dewaterers,  or  classifiers,  from  which  it  gravitates  under 
ground,  for  stope  filling.  The  destruction  of  the  cyanide  present  in  the 
residues  is  accomplished,  where  necessary,  by  the  addition  of  some  0.08  Ib. 
of  KMn(>4  per  ton  of  sand.  Since  its  introduction  on  the  Rand,  sand  fill- 
ing of  stopes  has  been  very  advantageously  adopted  in  other  countries 

Rand  Metallurgical  Costs. — In  this  practice,  the  following  comparison 
of  the  cost  per  ton  for  transferring  drained  sand  from  tank  to  tank  is :  (a)  • 


156  CYANIDING  OF  GOLD  ORES 

by  truck  haulage,  3.96  cents;  (b)  by  shoveling  and  conveyor  belt,  3.36 
cents;  (c)  by  shuttle  belts  and  main  conveyor  belts,  2.44  cents. 

In  offering  sand-leaching  cost-data,  it  may  be  well  to  cite  the  working  of 
a  few  representative  plants.  Again,  it  must  be  noted  that  no  comparisons 
are  here  intended,  local  conditions  being  at  too  great  a  variance.  Fifteen 
Rand  plants  show  that  an  average  of  57.6  per  cent  of  the  mill  pulp  is  leached 
as  sand  by  double  treatment,  the  proportion  ranging  from  70  per  cent  sand 
in  the  older  plants,  to  40  per  cent  sand  in  recent  plants.  The  economical 
limit  of  fine  grinding  of  the  sand  approximates  70  per  cent  through  100- 
mesh  (0.0058-in.  aperture).  With  intermittent  collecting,  revolving  dis- 
tributors are  preferred  to  hose-filling.  The  transfer  of  the  drained  sand 
to  the  treatment  tanks  is  by  truck-haulage,  or  by  shoveling,  in  cases 
where  the  collectors  or  settling  tanks  are  set  above  the  treatment 
tanks.  With  continuous  sand-filter  collecting,  the  collected  sand  is  trans- 
ferred by  conveyors  or  pumped  to  revolving  distributors.  The  residue 
is  discharged  by  belts,  trucks,  buckets,  or  sluicing  into  cone-dewaterers  for 
sand  filling  of  stopes.  The  total  ratio  of  solution  to  sand  for  dissolving 
and  washing  approximates  1.5  to  1.  About  four  days  are  allowed  for  col- 
lecting and  draining  and  seven  to  eight  days  for  treatment.  Approxi- 
mately one  ton  of  solution  is  precipitated  per  ton  of  sand.  The  strong 
solution  is  about  2  Ib.  KCN,  while  the  weak  is  about  0.5  Ib.  KCN. 

Figures  from  ten  Rand  plants  for  1912,  with  daily  sand  tonnage  running 
as  high  as  3500  tons,  show  the  following  averages : 

Total  cost  of  sand  treatment  per  ton  of  sand  (including  classification,  col- 
lecting, treating,  precipitating,  refining  and  residue  disposal),  ranging 

from  37  to  45  cents,  average 39.8  cents 

Average  extraction  on  $3.17  sand  heads 79.8  per  cent 

Average  extraction  in  recent  sand  plants 89.5  per  cent 

Costs  of  Rand  Plants. — The  capital  expenditures  for  various  capacities 
of  plants,  per  ton  of  sand  treated  per  twenty-four  hours,  are  as  follows: 
350  tons  of  sand  per  day,  $264;  700  tons,  $254;  1400  tons,  $244;  2800 
tons,  $233. 

Estimated  capital  costs  of  sand  plants  using  continuous  sand-filter 
collecting,  yielding  55  per  cent  sand,  are  given  as  follows:  275  tons  of  sand 
per  day,  $230  per  ton  of  daily  sand  capacity;  550  tons  of  sand,  $212; 
1100  tons  of  sand,  $195;  2200  tons  of  sand,  $177. 

Cost  of  Homestake  Plant. — The  actual  combined  cost  of  the  two  Home- 
stake  sand  plants,  of  a  total  capacity  of  2500  tons,  amounted  to  $255  per 
ton  of  daily  sand  capacity  This  figure  includes  power,  heating,  pre- 
cipitating, and  clean-up  plants. 


SLIME  TREATMENT  157 


(2)     FILTER-SLIME  TREATMENT 

Decantation. — As  largely  practiced  on  the  Rand,  the  slime  is  settled 
in  slime  collectors,  large  cone-bottom  tanks  provided  with  outlets  for  the 
escape  of  the  water,  and  in  adjustable  decanter,  all  the  slimes  settling  to 
the  bottom  of  the  cone.  The  settled  slime,  containing  approximately 
50  per  cent  of  moisture,  is  then  sluiced  out  by  aid  of  a  weak  cyanide  solu- 
tion and  is  pumped  to  the  "  first  settlement  tank."  Here  the  charge  is 
kept  in  agitation  by  means  of  a  circulating  pump  which  withdraws  from 
the  bottom  of  the  tank  to  discharge  at  the  top.  The  charge  when  suf- 
ficiently agitated  is  allowed  to  settle  and  the  gold-bearing  solution  decanted, 
then  precipitated  at  the  zinc  boxes.  The  settled  charge  of  slime  in  this 
tank  is  transferred  to  the  second  settling  tank  with  addition  of  weak  cyanide 
solution,  and  after  settling  and  decanting  of  the  supernatant  solution,  the 
thickened  slime  is  discharged,  using  for  that  purpose  sufficient  fresh  water. 
Settlement  is  aided  by  the  judicious  use  of  lime.  This  method,  while  sim- 
ple and  inexpensive,  has  the  drawback  that  there  is  an  inevitable  loss 
through  the  imperfect  washing,  though  for  low-grade  ores  such  loss  is 
inconsiderable. 

All-sliming. — Except  in  the  case  of  certain  low-grade  mines,  such  as  the 
Rand  and  Homestake,  the  earlier  practice  of  comparatively  coarse  grind- 
ing, then  treating  the  sand  and  slime  separately  has  largely  given  way  to 
the  practice  of  finer  grinding  and  treating  the  entire  pulp.  "  All  sliming  " 
is  the  term  applied  to  this.  Where  a  separate  sand  slime  treatment  has 
prevailed  the  two  products  have  been  separated  by  classifier,  the  slime  of 
such  fineness  that  90  per  cent  of  it  will  pass  a  200-mesh  screen.  In  the  all- 
sliming  plants  the  pulp  varies  from  60  per  cent  to  90  per  cent  through  200- 
mesh,  depending  upon  the  economic  limit  of  fine  grinding  and  subsequent 
treatment.  In  the  sand-slime  plant  the  grinding  is  done  in  water,  while 
where  all-sliming  prevails,  this  is  done  in  cyanide  solution. 

(3)     SLIME  AGITATION 

This  is  done  for  the  purpose  of  bringing  the  slimed  ore  intimately  in 
contact  with  the  cyanide  solution  and  with  air  according  to  Eisner's 
reaction. 

The  matter  of  proper  dilution  at  which  pulp  should  be  agitated  is  one 
demanding  careful  consideration  for  each  individual  plant.  Experimental 
work  points  to  high  dilutions,  but  practical  considerations  generally  neces- 
sitate as  thick  a  pulp  as  can  be  regularly  drawn  from  the  thickeners.  A 
dilution  of  1.5  to  1  is,  perhaps,  an  average.  Inasmuch  as. the  cost  of 
agitating  a  tori  of  slime  for  twenty-four  hours  should  not  exceed  3  cents,  it 
is  evident  that  the  agitation  period,  to  be  commensurate  with  this  lovv  cost, 


158 


CYANIDING  OF  GOLD  ORES 


should  be  well  advanced  toward  the  point  where  further  dissolution  ceases. 
In  determining  the  proper  length  of  agitation,  such  items  as  capital  charge 
on  the  agitators  and  accessories,  consumption  of  cyanide,  lime,  and  other 
chemicals,  possible  premature  precipitation  of  dissolved  metals  in  pro- 
longed agitation,  and  the  mechanical  cost  of  agitating,  should  be  carefully 
balanced  against  the  net  returns  from  the  metals  dissolved.  Then,  too, 
the  possibility  of  decreasing  the  fineness  of  grinding  by  prolonging  the  agi- 
tation should  be  examined. 

The  modern  trend  is  in  favor  of  continuous  agitation,  as  compared  with 
the  former  intermittent  system.  Change  of  solution  during  agitation  is 
found  advisable  with  certain  ores;  although  with  present-day  agitators, 
capable  of  giving  ample  aeration,  the  entire  agitation  step  is  frequently 
carried  out  in  one  stage. 

The  types  of  agitators  may  be  divided  into  (1)  pneumatic,  (2)  mechani- 
cal, (3)  combined  mechanical  and  pneumatic,  (4)  pump  transfer  system. 

(1)     PNEUMATIC  AGITATORS 

These  include  the  Parral  and  the  Pachuca 
(or  Brown) .  The  latter  is  an  efficient  machine 
for  concentrates,  but  in  the  majority  of  cases 
the  cost  of  pumping  the  pulp  into  these  tall 
tanks  is  one  serious  thing  against  them.  It  is 
also  costly  to  erect. 

The  Pachuca  or  Brown  tank,  shown  in  Fig. 
100,  consists  of  a  steel  cylinder  ordinarily  30  ft. 
high  and  10  ft.  in  diameter,  terminating  below 
in  a  cone  having  an  angle  of  60°  at  the  vertex. 
Vats  60  ft.  high  and  15  ft.  in  diameter  are  being 
constructed  in  some  mills.  In  the  lower  part 
of  the  cone  there  is  a  6-in.  pipe  with  a  gate 
valve  for  discharging  the  contents  of  the  vat 
after  treatment.  The  apparatus  in  the  interior 
of  the  vat  consists  of:  (1)  A  central  10-in.  tube 
called  the  elevator.  (2)  A  pipe  for  compressed 
air  passes  through  the  center  of  the*  elevator 
and  rests  on  the  bottom  of  the  cone.  This 
pipe  is  closed  at  the  bottom  end,  but  has  a 
number  of  small  holes  in  it  at  the  level  of  the 
bottom  of  the  elevator  tube.  These  holes  are 
covered  by  a  section  of  rubber  hose  slipped 

FIG.  100.— Pachuca  Tank.       over  *^e  P*Pe  an(^  having  its  lower  end  tied  to 

the  pipe  with  wire  just  below  the  holes  in  the 
pipe,  so  that  the  hose  forms  a  sort  of  collar  valve  through  which  the  air 


THE  PACHUCA  TANK  159 

may  escape  into  the  vat,  while  the  pulp  is  prevented  from  entering  the 
pipe.  When  the  pressure  of  air  in  the  pipe  is  greater  than  that  due  to  the 
column  of  slime  hi  the  vat  the  air  escapes  through  the  collar  valve  and 
passes  up  through  the  end,  but  has  a  number  of  small  holes  in  it  at  the 
level  of  the  bottom  elevator  tube,  carrying  the  slime  with  it.  (3)  Another 
air  pipe  with  its  collar  valve  is  placed  outside  of  the  elevator  to  keep  the 
slime  in  circulation  during  the  filling  and  emptying  of  the  vat.  (4)  An 
adjustable  agitating  device  consisting  of  an  annular  pipe  having  several 
small  pipes  with  collar  valves  so  arranged  that  compressed  air,  water,  or 
solution  may  be  forced  through  them  in  order  to  wash  off  any  sand  or 
slime  which  may  have  deposited  on  the  sides  of  the  cone  after  the  agita- 
tion has  been  stopped. 

The  method  of  operation  to  cause  the  agitation  or  circulation  of  the 
slime  in  these  vats  is  very  simple,  being  as  follows:  When  the  vat  is  filled 
with  pulp  and  solution,  the  valve  is  opened,  admitting  compressed  air  into 
the  bottom  of  the  elevator  where  it  mixes  with  the  pulp  in  the  elevator, 
and  as  this  mixture  is  lighter  than  the  pulp  in  the  vat,  it  rises  to  the  top  of 
the  elevator  and  overflows  into  the  vat,  while  another  portion  of  the  slime 
in  the  vat  enters  the  elevator  at  the  bottom  which  in  turn  is  raised  to  the 
top  and  overflows,  thus  obtaining  a  perfect  and  continuous  circulation  of 
the  pulp  as  long  as  the  compressed  air  is  allowed  to  enter.  Upon  starting 
the  circulation  the  air  pressure  must  be  greater  than  that  of  the  column  of 
slime,  but  as  soon  as  the  circulation  is  well  established  less  pressure  is 
required,  it  having  been  found  in  practice  that  while  50  Ib.  pressure  is 
required  to  start  the  circulation,  as  soon  as  the  sand  and  slime  which  have 
settled  in  the  bottom  of  the  cone  have  been  cleared  out  by  the  scouring  of 
the  circulating  pulp,  the  circulation  may  be  maintained  by  a  pressure  of 
but  25  Ib.  per  square  inch.  The  quantity  of  air  required  in  any  particular 
case  depends  on  the  proportion  of  sand  to  slime,  the  fineness  of  the  pulp, 
and  the  viscosity  of  the  slime.  Ordinarily,  in  the  Pachuca  plants,  100  cu.  ft. 
per  minute  is  used  to  maintain  a  vat  containing  100  tons  of  slime  in  active 
circulation  and  to  prevent  the  settlement  of  the  sand  on  the  sides  of  the 
cone  bottom,  but  in  the  Goldfield  Consolidated  plant,  Nevada,  and  at  the 
Komata  Reefs,  New  Zealand,  only  30  cu.  ft.  per  minute  is  used. 

(2)     MECHANICAL  AGITATORS 

Mechanical  Agitators. — For  the  agitation  of  heavy  sulphides  we  have 
the  old  mechanical  agitator  with  plowshoes  revolving  at  16  to  18  R.P.M. 
Much  power  is  required  to  operate  it. 

Top-drive  paddle-arm  stirrers  or  agitators  were  used  from  the  start  of 
the  slime-agitation  process,  and  with  proper  design,  still  hold  a  fair  place 
at  this  day.  The  bottom-drive  agitator,  as  used  in  Kalgoorlie,  Mexico, 


160  CYANIDING  OF  GOLD  ORES 

and  the  Rand,  eliminates  the  submerged  step  bearing  of  the  top  drive, 
allows  of  a  simpler  overhead  construction,  and  due  to  the  increased  rigidity 
of  the  drive,  operates  with  very  moderate  power.  Assuming  a  tank  30  ft. 
diameter  by  12  ft.  deep  as  an  approximate  standard,  a  satisfactory  agitating 
speed  for  the  bottom  drive  type  may  be  given  as  8  R.P.M.  and  the  power 
as  6  H.P.  In  addition,  the  power  for  furnishing  20  cu.  ft.  of  free  air  per 
minute  at  10  Ib.  pressure,  for  requisite  aeration,  amounts  to  1  H,P.  This 
type  of  agitator  works  quite  satisfactorily,  although  it  is  troublesome  to 
start  up  after  a  protracted  shut-down,  tends  to  bank  up  near  the  center 
and  gives  en  masse  agitation.  Inclined  baffle-boards  bolted  to  the  side  of 
the  tank  will  materially  assist  agitation. 

The  average  erected  cost  of  a  30Xl2-ft.  bottom-drive,  steel-tank, 
mechanical  agitator  may  be  given  as  $2395.  Such  a  tank  will  have  an 
available  depth  of  11  ft.  6  in.,  and,  with  pulp  at  a  dilution  of  1.5,  will  hold 
134  tons  of  dry  slime,  making  the  total  erected  cost  $17.87  per  ton  of  dry 
slime  capacity. 

Agitation  of  the  Slime. — By  fine  grinding  even  microscopic  particles 
of  gold  have  been  set  free  from  the  gangue  or  waste.  These  particles  can 
be  dissolved  provided  they  can  be  brought  intimately  in  contact  with  the 
cyanide  solution  in  presence  of  air.  This  is  done  by  agitation,  preferably 
using  an  air  jet  for  so  doing.  In  this  way  the  solution  is  well  circulated 
and  is  brought  in  contact  with  each  particle  of  the  slimed  material  as 
ordinary  stirring  or  mixing  will  not  do.  Even  then  the  solution  of  the  gold 
particles  takes  hours  of  time,  and  agitation  is  continued  until  assay  shows 
that  further  dissolution  has  ceased. 

(3)     COMBINED  MECHANICAL  AND  PNEUMATIC  AGITATORS 

These  include  the  Dorr,  the  Trent,  and  the  Hendryx  machines.  The 
Dorr  agitator  appears  to  be  the  cheapest  in  first  cost,  requires  less  power, 
and  will  operate  with  the  least  amount  of  trouble.  It  is  well  suited  for  a 
concentrate,  ground  to  pass  a  200-mesh  screen.  The  Trent  and  Hendryx 
machines  are  not  well  suited  for  heavy  sulphide  material. 

Dorr  Agitator.  This  comparatively  new  type  of  agitator,  combining 
mechanical  and  pneumatic  agitation,  has  met  with  merited  success.  It  is 
usually  operated  at  3  R.P.M.,  the  speed  depending,  of  course,  on  the  char- 
acter of  the  pulp.  Under  average  conditions,  a  30Xl2-ft.  Dorr  agitator 
will  require  about  1.5  H.P.  for  moving  the  arms,  at  3  R.P.M.,  and  2.5  H.P. 
for  furnishing  30  cu.  ft.  of  free  air  per  minute  at  20  Ib.  pressure,  a  total  of 
4  H.P.  The  total  erected  cost  of  such  an  agitator  with  steel  tank  will 
approximate  $2510.  With  an  available  depth  of  11  ft.  6  in.  and  with  pulp 
at  a  dilution  of  1.5,  this  agitator  will  hold  134  tons  of  dry  slime,  making  the 
total  erected  cost  $18.73  per  ton  of  dry  slime  capacity. 


THE  DORR  AGITATOR  161 

Fig.  101  is  a  view  of  this  tank.  In  a  flat-bottom  steel  tank  is  a  central 
vertical  cylinder  or  pipe  carried  by  a  shaft  supported  from  the  top  of  the 
tank  and  having  two  stirring  arms  with  plows  as  in  the  Dorr  thickener. 
The  plows  both  agitate  the  pulp  and  draw  it  toward  the  center.  The  pulp 
is  raised  through  the  cylinder  by  means  of  air,  supplied  under  pressure  from 
an  air  pipe  whose  nozzle  points  upward  at  the  foot  of  the  vertical  cylinder. 
The  pulp,  rising  through  the  cylinder,  is  delivered  by  two  opposite  launders 
attached  to  the  cylinder  by  which  it  is  distributed  over  the  surface  of  the 
liquid  contents.  A  continuous  supply  of  fresh  pulp  enters  at  the  intake 
at  the  left,  and  is  distributed  through  the  tank,  and  after  thorough  agita- 


FIG.  101.— The  Dorr  Agitator. 

tion  and  aeration,  escapes  by  the  outflow  on  the  right.  Thus  we  have  a 
method  of  continuous  agitation  which  has  its  advantages  over  agitating  in 
charges. 

AGITATION  TREATMENT 

Dilution  or  ratio  of  solution  to  dry  ore  in  the  pulp  undergoing  treat- 
ment is  generally  recognized  as  a  factor  to  which  proper  attention  must  be 
paid  if  the  highest  extraction  is  to  result.  If  a  higher  dilution  is  used  than 
is  necessary  the  capacity  of  the  plant  is  cut  down,  and,  on  the  other  hand,  if 
the  dilution  be  too  low  the  maximum  extraction  is  not  attained.  It  is 
quite  impossible  to  give  general  figures  for  minimum  dilution  as  these 
figures  vary  for  different  ores.  In  general,  lower  dilutions  can  be  used  with 
gold  ores  than  with  silver  ores,  probably  on  account  of  the  greater  weight 
of  metal  to  be  dissolved  in  the  latter  case.  Other  conditions  being  equal,  it 
appears  that  lower  dilutions  can  be  used  with  the  Pachuca  agitator  than 


162  CYANIDING  OF  GOLD  ORES 

with  the  mechanical  agitator  on  account  of  the  greater  proportion  of  air 
brought  in  intimate  contact  with  the  pulp. 

The  time  of  treatment  depends  upon  the  character  of  the  silver  and  gold 
minerals  and  upon  their  degree  of  comminution,  and,  as  previously  pointed 
out,  also  upon  the  concentration  or  strength  of  the  solution  in  cyanide.  It 
therefore  follows  that  finer  grinding,  or  increase  of  the  cyanide  concentra- 
tion of  the  solution,  or  both,  will  in  general  result  in  reducing  the  time  of 
treatment  necessary.  In  cases  where  the  cost  of  power  is  high,  and  as  a 
consequence  the  cost  of  fine  grinding  would  be  excessive,  the  ore  may  be 
ground  to  the  point  where  the  minerals  are  liberated  from  the  gangue  and 
then  separated  into  sand  and  slime,  the  slime  being  treated  by  agitation 
followed  by  either  decantation  or  filtration.  The  sand  containing  the  coarse 
mineral  particles  requiring  a  long  period  of  contact  for  dissolution  can  be 
given  the  long  period  of  contact  necessary  at  a  reasonable  cost  by  leaching. 
Cases  of  this  kind  clearly  indicate  where  combined  sand  and  slime  treat- 
ment can  be  employed  to  advantage. 

Thickening  the  Slime. — As  a  preliminary  adjunct  to  any  filter  operation, 
the  pulp  is  first  settled  to  a  minimum  moisture  content;  in  average  practice 
this  will  be  found  to  approximate  closely  to  50  per  cent,  or  equal  parts  of 
liquid  and  solid.  In  some  plants  this  thickening  is  all  done  prior  to  dissolu- 
tion, while  in  others,  thickeners  are  used,  both  before  and  after  the  agitation. 

For  filtration,  the  moisture  should  be  reduced  to  such  a  point  that  the 
heavier  particles  will  remain  in  suspension,  or  at  least  settle  very  slowly 
during  the  cake-forming  period.  This  governs  the  uniformity  of  the  cake 
and  is  one  of  the  most  vital  points  in  all  filter  operations,  and  since  the 
size  of  the  largest  particles  is,  in  turn,  governed  by  the  limits  of  economic 
grinding,  the  required  buoyancy  is  best  obtained  by  proper  thickening. 

The  use  of  the  Dorr  continuous  thickener  has  become  almost  universal 
for  this  work  in  America,  and  is  being  largely  adopted  in  foreign  countries 
as  well.  The  machine  requires  a  minimum  of  power  and  attendance,  and 
when  used  in  conjunction  with  a  diaphragm  pump  or  air  lift,  for  elevating, 
the  discharge  may  be  operated  with  practically  no  loss  of  mill  head.  Dis- 
charges as  low  as  33  per  cent  moisture  are  obtained  in  the  Porcupine  district, 
but  on  other  ores  careful  attention  may  be  needed  to  obtain  60  fler  cent. 

On  the  Rand,  with  colored  labor  at  $0.75  or  less  per  day,  and  power  at 
$5.50  per  H.P.-month,  it  is  still  found  economical  to  retain  the  large  inter- 
mittent settlers.  These  are  steel  tanks,  varying  from  50  ft.  to  70  ft.  in 
diameter,  with  from  10-ft.  to  14-ft.  sides  and  cone  bottoms,  giving  an  addi- 
tional depth  of  from  4  to  8  ft.  Peripheral  overflows  and  adjustable  decant- 
ing arms  are  provided,  and  the  tanks  are  emptied  by  sluicing  the  settled 
slime  into  the  suction  of  a  centrifugal  transfer  pump. 

This  system  of  settling  or  dewatering  is  particularly  well  adapted  to 
African  conditions,  where  flat  open  mill  sites  are  used,  and  where  all  tanks 


AGITATION  TREATMENT  163 

are  unhoused.  But  it  is  to  be  noted  that  while  the  intermittent  settlers 
on  the  Rand,  particularly  during  the  warm  summer  months,  frequently 
settle  down  to  40  per  cent,  and  even  to  38  per  cent,  moisture,  the  best 
that  continuous  thickeners  seem  able  to  do  is  50  per  cent.  Since  the  extra 
10  per  cent  of  water  must  be  brought  up  to  treatment  strength  in  cyanide, 
and  then  wasted,  its  elimination  is  highly  desirable.  With  any  settling 
equipment  satisfactory  moisture  figures  are  seldom  attained  during  the 
winter  months,  when  60  per  cent  is  more  nearly  the  average  figure,  with,  of 
course,  the  higher  losses  in  cyanide  and  in  gold. 

The  area  required  for  proper  continuous  setting  is,  of  course,  mainly 
a  function  of  the  nature  of  the  ore  and  the  dilution  of  the  pulp  and  varies 
from  4  to  15  sq.  ft.  per  dry  ton  of  daily  capacity  with  a  pulp  feed  of  from 
90  per  cent  to  75  per  cent  moisture. 

With  intermittent  settling  as  practiced  on  the  Rand,  the  period  required 
for  discharging  introduces  an  additional  time  factor,  and  it  is  usual  to  allow 
14  to  25  sq.  ft.  per  ton  of  dry  slime.  It  should  be  noted  that,  in  this  prac- 
tice, practically  all  the  water  used  in  crushing  and  classification  goes  to  the 
slime  collectors,  which  thus  handle  a  feed  containing  from  90  to  95  per 
cent  moisture. 

The  cost  of  thickening  operations  will  vary  from  $0.005  to  $0.02  per 
ton  milled,  depending  upon  local  conditions  and  the  scale  of  operations. 

Naturally,  there  is  no  standard  size  recommended  or  used,  as  this 
depends  on  the  capacity  desired,  settling  qualities  of  the  pulp,  density  to 
which  thickening  is  to  be  carried,  clearness  of  overflow,  alkalinity,  tempera- 
ture, and  dilution  of  the  feed.  As  a  general  rule,  it  may  be  stated  that 
approximately  6  sq.  ft.  of  tank  area  are  needed  per  ton  of  granular  slime 
per  twenty-four  hours,  while  10  to  15  sq.  ft.  should  be  allowed  for  flocculent 
slime.  Several  installations  in  different  localities  show  an  average  of 
$2500  for  the  complete  erected  cost  of  a  standard,  steel,  30X12  ft.  unit. 
In  terms  of  tonnage  of  dry  slime  handled  per  twenty-four  hours,  the  capac- 
ity of  a  30X20-ft.  standard  thickener,  under  normal  conditions,  may  be 
given  as  125  tons  of  granular  slime  and  65  tons  of  flocculent  slime,  making 
the  erected  cost  of  the  thickener  $20  and  $38.46,  respectively,  per  ton  of 
daily  capacity. 

The  Dorr  Continuous  Thickener. — As  shown  in  Fig.  102,  this  30  X  12-ft. 
tank  has  a  slowly  moving  central  vertical  shaft  with  radial  arms  equipped 
with  plows  to  bring  the  thickened  settled  material  to  its  discharge  point 
at  the  center.  The  thick  slime  discharge  is  pumped  to  another  tank  for 
further  treatment.  The  feed  launder  delivers  to  a  central  drop  pipe  so  as 
to  cause  no  agitation.  The  clear  solution  escapes  to  the  peripheral  launder 
of  the  tank  and  thence  overflows.  The  vertical  shaft  will  revolve  about 
once  hi  twelve  minutes. 

Fig.  103,  a  sectional  elevation  of  this  thickener,  shows  its  operation. 


164 


CYANIDING  OF  GOLD  ORES 


The  feed  of  this  pulp  at  the  center  of  the  tank,  the  overflow  of  clear  liquid 
at  the  periphery  and  the  discharge  of  thickened  pulp  are  continuous. 
There  are  four  zones  of  settlement.  At  the  top  is  a  zone  of  clear  water  A, 
beneath  this  is  zone  B,  consisting  of  flocculated  pulp  of  uniform  consistency; 


FIG.  102. — The  Dorr  Continuous  Thickener. 


FIG.  103. — Dorr  Thickener,  Showing  Slime-settling  Zones. 

directly  beneath  this  is  a  transition  zone  C,  and  at  the  bottom  a  zone  of 
pulp  which  is  undergoing  compression.  The  pulp  in  zones  B  and  C  is 
termed  "  free  settling." 

CONTINUOUS  COUNTER-CURRENT  DECANTATION 

This  method  of  separating  dissolved  values  from  treated  slime,  by  means 
of  a  series  of  Dorr  continuous  thickeners,  is  becoming  very  popular  in 


COUNTER-CURRENT  DECANTATION  165 

America,  particularly  in  small  plants,  and  more  especially  in  those  treating 
a  granular  product,  which  readily  settles  to  40  per  cent  moisture  or  less 
and  is  amenable  to  treatment  with  very  low  cyanide  strengths. 

In  operation,  the  slime  passes  through  a  series  of  tanks,  the  thick  under- 
flow of  each  being  diluted  with  solution  overflowing  the  second  following 
thickener  of  the  series.  The  solids  thus  move  constantly  in  one  direction, 
while  the  solutions  travel  in  the  opposite  direction.  Water,  in  quantity 
sufficient  to  replace  the  moisture  finally  discharged  with  the  tailings,  is 
added  at  the  final  thickener.  The  solution,  traveling  successively  toward 
the  head  of  the  series  and  mixing  with  constantly  richer  pulp,  is  finally 
used  in  the  crushing  department,  whence  it  overflows  the  first,  or  primary, 
settler  and  is  sent  to  precipitation.  The  tanks  are  generally  set  so  that 
solutions  gravitate  throughout  the  series,  and  the  necessary  elevation  of 
the  thick  underflow  is  made  either  with  diaphragm  pumps  or  with  air  lifts. 

In  most  plants  where  continuous  decantation  has  been  successfully 
used,  underflows  of  35  to  40  per  cent  moisture  are  usually  maintained  and 
solution  equal  to  from  four  to  six  times  the  weight  of  the  ore  is  clarified 
and  precipitated.  Under  these  conditions,  the  recovery  of  dissolved  metals 
is  excellent,  but  the  loss  of  cyanide  is  higher  than  with  filters. 

This  mechanical  loss  of  cyanide  is  recognized  as  one  of  the  principal 
factors  limiting  the  use  of  continuous  decantation  without  filters,  where 
even  moderately  strong  solutions  are  used.  Solutions  of  less  than  \  Ib. 
per  ton  KCy  are  seldom  precipitated  hi  American  practice,  and  at  this 
strength,  the  mechanical  loss  in  the  final  residue  will  vary  from  J  Ib.  to  as 
high  as  1  Ib.,  depending  on  underflow  moistures.  In  milling  silver  ores, 
there  will  be  an  additional  loss,  owing  to  the  fact  that  dissolution  of  the 
metal  continues  as  long  as  the  slime  is  in  contact  with  solution. 

The  actual  operating  cost  of  the  thickeners  is  very  low,  and  the  sim- 
plicity of  the  plant,  enabling  labor  and  supervision  to  be  reduced  to  a  min- 
imum, must  appeal  strongly  to  operators  of  small  plants. 

As  a  most  interesting  adaptation  of  continuous  decantation  and  vacuum 
filtration  may  be  cited  the  practice  at  the  recently  enlarged  Hollinger  mill 
at  Porcupine.  Small  thickeners,  operating  as  classifiers,  separate  the  tube- 
mill  product  into  amorphous  and  granular.  The  former  is  thickened  and 
sent  to  a  vacuum  filter  for  washing,  without  other  agitation  than  that 
obtained  in  grinding,  pumping  and  thickening. 

The  granular  portion  is  concentrated,  thickened,  agitated,  and  then 
sent  to  counter-current  decantation  tanks  for  washing.  This  system 
eliminates  amorphous  or  colloidal  material  from  the  bulk  of  the  tonnage, 
where  a  very  thick  underflow  is  imperative,  and,  at  the  same  time,  fur- 
nishes a  satisfactory  product  for  the  filter. 

This  is  now  known  as  the  C.  O.  D.  system.  The  flow  sheet  of  it,  shown 
in  Fig.  104,  illustrates  the  method  where  there  are  four  Dorr  thickeners. 


166 


CYANIDING  OF  GOLD  ORES 


W,  X,  Y  and  Z  in  series.  It  is  assumed  that  the  crushing  is  done  in  cyanide 
solution,  the  overflow  from  the  thickening  Tank  X,  being  returned  to  the 
mill  for  mill  solution.  The  ground  pulp  enters  tank  W,  whose  overflow,  called 
the  pregnant  solution,  goes  to  the  next  step  in  cyaniding,  the  precipitating 
of  the  gold  from  the  solution.  After  having  here  deposited  its  gold  con- 
tent the  solution,  now  called  "  barren,"  is  used  to  dilute  the  underflow 
of  thickener  X,  as  it  enters  tank  F.  The  overflow  of  thickener  Z  is  also 
mixed  into  the  feed  to  F,  receiving  also  water  for  washing  its  pulp,  which  is 
then  sent  to  waste.  The  overflow  from  F  meantime  enters  X}  together 
with  the  partially  exhausted  thickened  pulp  from  W.  It  is  thus  seen 
that  pulp  passes  from  tank  to  tank  from  left  to  right,  losing  more  and  more 
of  its  gold,  to  be  discharged  and  exhausted  from  Z;  while  the  wash-water 

flowing  into  Z  picks  up  an  increasing  load 
of  gold  as  it  encounters  the  progressively 
richer  pulp  in  its  passage  to  the  left.  It  fi- 
nally leaves  W  at  its  full  possible  strength. 
This  method  is  best  adapted  to  low- 
grade,  easily  leached  and  settled  ores,  and 
it  dispenses  with  agitation  other  than  that 
resulting  from  the  crushing  and  classify- 
ing and  the  movement  through  the 
thickeners,  and  likewise  with  filtration 
about  to  be  described. 

At  the  Hollinger  mill,  Porcupine  Dis- 
trict, Ontario,  Canada,  where  some  50,000 
tons  of  a  soft  quartz  ore  with  schist  and 
pyrite  is  treated  monthly,  there  are  five 
sets  of  40-ft.  tanks  as  just  described.  The 
.^zinc-dust  Feeder  tanks  are  arranged  with  a  difference  of 
elevation  of  2J  ft.  between  the  steps,  the 
last  of  the  series  Z,  being  the  highest  so 
the  solution  flows  from  tank  to  tank  and 
to  precipitation.  Thence  it  is  pumped  to 
the  mill-solution  feed-tank.  Th<i  cost  of 
decanting  is  given  at  2.09  cents  and  the 
tailings  carry  9.75  cents  per  ton. 


gj'f   precipitate  Press 

a 


Oil-fired  Tiltinog  Furnace 
ullion 


FIG.  104.  —  Continuous  Counter-Cur- 
rent Decantation. 


FILTRATION    OR    SEPARATION    OF    METAL-BEARING    SOLUTION    FROM 

SLIME  RESIDUE 

Probably  the  most  vital  point  in  the  practical  application  of  the  cyanide 
process  is  the  filtration  or  separation,  after  dissolution,  of  the  metal- 
bearing  solution  from  the  slime  residue.  Certainly  no  other  point  has 
called  forth  such  a  combination  of  inventive  ingenuity  and  practical  ability 


SLIME  FILTRATION  167 

as  has  been  expended  in  developing  a  satisfactory  technical  and  economical 
solution  of  this  problem. 

The  reason  for  this  is  readily  appreciated  when  one  stops  to  consider 
that  for  every  unit  of  dissolved  metal  finally  discharged  with  the  residue, 
there  is  incurred  a  net  loss  equal  to  the  market  value  of  the  unit  plus  the 
mechanical  loss  of  cyanide,  amounting  to  a  further  $0.05  to  $0.10  per  ton. 
Since  filtration  or  decantation  is  almost  the  last  step  in  gold  and  silver  pro- 
duction, and  amounts  at  most  to  5  per  cent  of  the  total  cost,  there  is 
obviously  every  incentive  to  obtain  the  highest  possible  efficiency. 

The  amount  and  character  of  this  material  produced  will  depend 
chiefly  upon  the  nature  of  the  ore  and  upon  the  degree  of  comminution 
necessary  to  obtain  an  economic  extraction.  Many  so-called  "  all-slime  " 
plants  find  it  feasible  to  agitate  and  filter,  or  decant,  a  product  of  which 
fully  40  per  cent  will  remain  upon  a  200-mesh  screen,  while  others  grind  to  a 
point  where  only  a  fraction  of  1  per  cent  will  remain  upon  this  mesh.  The 
governing  factors  are  purely  individual,  such  as  size  and  nature  of  plant, 
location,  character  of  ore,  etc.,  and  can  hardly  be  generalized. 

All  ores  after  fine  grinding  may  be  classified  into  two  products — granular 
and  amorphous,  and  most  of  the  difficulties  experienced  in  slime  filtration 
and  decantation  may  be  traced  to  extreme  conditions  as  regards  either  the 
one  or  the  other. 

An  undue  amount  of  coarse,  granular  material  will  not  only  increase 
the  power  consumption  in  the  agitators  and  thickeners,  but  will  also  choke 
and  cause  frequent  interruptions,  and  unless  filtered  at  50  per  cent  moisture, 
or  less,  will  result  in  classification  and  uneven  washing. 

On  the  other  hand,  where  the  percentage  of  the  amorphous  product 
is  high,  it  is  seldom  possible  to  thicken  to  less  than  60  per  cent  moisture, 
even  with  a  very  large  settling  area,  and  a  pulp  of  this  dilution  will  not  give 
economic  results  with  either  continuous  or  intermittent  decantation,  and 
even  for  filter  treatment  requires  a  largely  increased  area  and  frequently 
results  in  slow  and  imperfect  washing.  Cracks  and  channels  will  also 
develop  with  any  type  of  filter  in  which  the  cake  is  exposed  to  the  air 
before  or  during  washing.  . 

It  may  be  safely  stated  that  for  ideal  slime-washing  results  with  any 
of  the  methods  now  in  vogue,  an  all-slime  product  should  fulfill  the  following 
conditions. 

It  should  be  ground  in  closed  circuit  with  suitable  classifiers  so  that  not 
more  than  25  per  cent  of  the  total  product  will  remain  upon  200-mesh 
screen  and  all  of  the  metallic  or  sulphide  portion  will  pass  the  same  aperture. 

It  should  settle  from  85  per  cent  moisture  to  at  least  50  per  cent  with 
a  continuous  settling  area  of  6  to  10  sq.  ft.  per  ton  of  dry  solids  per  day. 

The  various  processes  and  machines  now  in  use  include  95  per  cent  of 
the  slime  tonnage  produced  in  the  cyanide  process  throughout  the  world. 


168 


CYANIDING  OF  GOLD  ORES 


MAIN  SYSTEMS   OF  FILTERING 

1.  Thickening. — The  Dorr  continuous  filter;  settling  tanks. 

2.  Vacuum  Filtration. — Butters,  leaf ;   Moore,  Oliver,  drum;   Portland, 
drum;   Bidgway,  leaf;  American,  continuous  suction. 

3.  Pressure  Filtration. — Merrill  sluicing  plate-and-frame  press;  Dehne, 
plate-and-frame-press;    Kelly,   enclosed-leaf;     Burt,    revolving   cylinder; 
Sweetland. 

4.  Continuous  Decantation. — Dorr  system. 

5.  Intermittent  Decantation.— Rand  system. 


FIG.  105.— The  Butter  Vacuum  Filter. 

(2)  VACUUM  FILTRATION 

The  Butters  Vacuum-leaf  Filter. — Fig.  106  is  an  elevation  and 
Fig.  107  a  view  of  a  filter  leaf  on  which  a  layer  of  sfcme  has 
been  built  up  and  part  of  it  removed  to  show  its  thickness.  A  vacuum 
filter  plant  consists  of  several  filter  tanks  (see  Fig.  105)  with  vertical 
sides  and  a  V-shaped  bottom,  in  which  are  suspended  a  number  of 
filter-leaves  constructed  of  pipe,  with  either  cocoanut-matting,  wooden 
laths,  or  ripple  iron  as  a  support  for  a  cloth  which  is  sewn  around  and  covers 
the  whole  as  shown  in  Fig.  106.  Leaves  vary  in  construction  and  5  by  9  ft. 
is  a  handy  size.  Connected  with  the  vats  is  suitable  piping  and  cen- 
trifugal pump  of  large  capacity  for  filling  and  emptying  with  slime,  wash- 
solution,  or  water  as  desired.  In  operation,  the  filter  vat  is  filled  with  pulp 


VACUUM  FILTRATION 


Wooden 
Cleat 


nnection  to 
Manifold 
"-Oregon  Beam 

C.I.  Shoe.  One  at 
each  End  of  Beam 


Duck  sewed  along  Top 

of  Pipe  and  Loose 

Ends  Clamped  between 

Beam 


Spacer 


ORIGINAL  FILTER  BEAM 


PRESENT  EILTER  BEAM 


to  a  point  over  the  top  of  the  filter  leaves,  and  a  valve  opened  connecting 
the  vacuum  pump  directly  with  the  filters.  Clear  cyanide  solution  is 
drawn  from  within  the  filter  leaves  and  delivered  to  a  clarifying  tank  for 
precipitation,  the  slime  remaining  as  a  cake  on  the  outside  of  the  filters. 
The  filters  are  kept  submerged  by  refilling  the  vat  at  intervals,  and  jets  of  air 
are  introduced  at  the  points  of  each  hopper  to  keep  the  slime  in  suspension. 
When  a  cake  1  in.  to  1J  in. 
thick  has  been  formed,  the  sur- 
plus pulp  is  pumped  back  to  the 
stock  pulp  vat,  and  the  box  is 
then  filled  with  solution  to  wash 
the  cake.  During  the  time  of 
forming  and  washing  the  cake, 
the  vacuum  is  maintained  at  the 
highest  possible  point,  but  when 
the  cake  is  exposed  to  the  air 
during  the  transfer  of  pulp  and 
wash  solution  the  vacuum  is  re- 
duced to  5  in.  to  prevent  the 
cake  cracking.  It  is  possible  to 
form  a'  cake  with  some  ores  in 
five  minutes,  while  others  take 
up  to  ninety  minutes. 

Sufficient  wash  solution,  about 
2  tons  per  ton  of  slime,  is  drawn 
through  the  cake  by  the  vacuum 
pump  to  effect  a  complete  dis- 
placement of  the  original  valu- 
able solution.  A  portion  of  this 
solution  goes  to  the  clarifying 
tank  before  precipitation.  When  the  wash  is  complete,  the  vacuum  is  dis- 
connected, and  a  flow  of  solution  or  air  is  introduced  into  the  interior  of  the 
filters,  causing  the  cakes  to  drop.  The  mass  of  thick  sludge  remaining  in 
the  vat  is  diluted  with  water  and  agitated  with  air  for  a  few  minutes  to 
make  a  homogeneous  pulp  of  1  to  1,  which  is  then  pumped  to  the  residue 
pond.  Typical  cycles  of  operation  are  as  follows: 


flLTER  FRAME  COMPLETE 


FIG.  106.— Butter's  Filter-frame. 


Rand  Ore. 

TonopahOre. 

Filling  vat  and  forming  cake  minutes                                     

45 

85 

Transferring  pulp  and  solution  and  wash 

70 

95 

Discharging                                 

20 

15 

Total   minutes             

135 

195 

170 


CYANIDING  OF  GOLD  ORES 


The  Oliver  continuous  revolving  filter  as  shown  in  Fig.  108  consists  of  a 
drum  or  cylinder  with  open  ends,  rotating  on  a  horizontal  axis,  with  the 
lower  portion  submerged  in  a  tank  containing  the  pulp  to  be  filtered.  The 
surface  of  the  drum  is  divided  into  compartments  or  sections,  the 


FIG.  107.— Slime-cake  partly  removed  to  show  section. 


FIG.  108.— Oliver  Filter. 


FIG.  109.— The  American  Filter. 


divisions  running  parallel  to  the  shaft.  These  sections  are  covered  by  a 
screen,  and  a  filter  medium  is  stretched  over  it,  being  held  in  place  and  pro- 
tected from  wear  by  a  wire  winding.  Each  section  of  the  drum  is  con- 
nected by  two  pipes  passing  through  a  hollow  trunnion  to  an  automatic 
valve  which  controls  the  application  of  the  vacuum  for  forming  and  wash- 


SUCTION  FILTRATION 


171 


ing  the  cake,  and  the  admission  of  air  for  its  discharge.  A  scraper  is  fitted 
across  the  tank  and  rests  on  the  wire  winding  so  that  the  washed  cake  is 
removed  when  released  by  air.  Vacuum  filtration  costs  vary  from  5  to  10 
cents  per  ton. 

The  American  Continuous  Suction  Filter. — Fig.  109  shows  a  view  of  a 
three-disk  filter  and  Fig.  110  an  end  view.  Each  disk  is  made  up  of  eight 
segments.  At  one  end  of  the  shaft  is  a  distributing  valve  having  eight 
openings,  each  to  one  of  the  eight  segments.  The  valve  housing  has  three 
inner  recessed  ports.  Port  No.  1  on  the  underside  connects  to  the  filtrate 
suction  line  and  applies  suction  to  the  four  submerged  leaf-segments. 
Port  No.  2  connects  with  the  wash-water  suction  and  drying  line,  taking 


Reyolytng  Filter  Leaf 


Leaf  Segment 
Baffle  Plates 


FIG.  110. — End  View  of  "American"  Filter. 

care  of  the  three  upper  left-hand  leaf  segments.  Port  No.  3  admits  com- 
pressed air  to  a  segment  as  it  passes  this  port.  The  air  is  admitted  for  a 
few  seconds  only,  inflating  the  filter  cloth  of  the  segment  and  loosening  the 
formed  cake  thereon,  while  the  scrapers  remove  it  so  that  the  product 
drops  into  a  hopper  or  conveyor  below. 

(3)     PRESSURE  FILTRATION 

Both  the  Kelly  and  the  Dehne  presses  have  been  extensively  used. 
They  have  the  advantage  over  the  suction  filters  that  high  pressure  can  be 
carried,  since  against  10  Ib.  suction  in  the  one  we  can  carry  50  Ib.  or  more  in 
the  pressure  filter.  This  makes  for  quick  filtering  and  washing. 

Dehne  presses  are  used  extensively* for  concentrate  in  Western  Australia 
and  at  Waihi,  New  Zealand.  Three-in.  cakes  are  formed  at  40-lb.  pressure 
and  are  washed  satisfactorily. 


172 


CYANIDING  OF  GOLD  ORES 


The  Kelly  Filter. — Fig.  Ill  shows  a  twin  unit  of  a  filter,  the  right-hand 
unit  open  for  discharging  the  washed  precipitate.      Into  the  closed  unit 

at  the  left  the  pulp  is  pumped 
filling  the  cylinder  around  the 
filter  leaves  as  seen  at  the  right- 
hand  unit.  The  escaping  solu- 
tion through  pipes  that  connect 
to  each  leaf  is  discharged  to  a 
launder.  This  continues  until  a 
thick  cake  has  been  built  up  on 
the  leaves.  The  excess  of  un- 
filtered  slime  is  run  out  and  fresh 
water  is  introduced  to  wash  the 
caked  accumulation.  To  dis- 
charge, the  front  head  of  the 


FIG.  112.— Single  Kelly 
Filter  Press. 

press  is  unlocked  and  the  head 
with  the  attached  carriage  which 
sustains  the  filter  leaves  is  run 
out,  as  shown  in  the  right-hand 
unit,  this  unit  by  the  same  move- 
ment being  closed  and  locked. 
The  cakes  are  dislodged  by  in- 
flating the  filter  leaves  with  air. 
The  carriage  is  run  into  the 
cylinder  and  the  head  again 
automatically  locked  in  place. 

The  Kelly  press  is  used  at 
the  Goldfield  Consolidated  and 
the  Alaska  Treadwell  with  satis- 
factory results. 

The  Rectangular  or  Dehne  Filter  Press. — This,  as  shown  in  Fig.  113, 
consists  of  a  heavy  frame  carrying  forty  filter  leaves  42  in.  square,  one 
standing  at  the  left  leaning  against  the  press  and  forty  filter  frames,  one  of 


PRESSURE  FILTRATION 


173 


FIG.  113. — Rectangular  or  Dehne  Filter  Press. 


j 


FIG.  114.— Sweetland  Filter  Press. 


PIG.  115.— Merrill  Filter-press  Installation. 


174 


CYANIDING  OF  GOLD  ORES 


FIG.  116.— Merrill  Press-frame. 


these  showing  at  the  right.  The  filter  leaves  are  covered  by  a  canvas  filter 
cloth,  one  on  each  side.  Fig.  116 A  shows  the  roughened  surface  of  a  leaf,  or 
plate,  the  surface  grooved  so  that  the  filtered  solution  is  carried  by  the 
grooves  to  the  outlet  channels,  circular  openings  at  the  sides  of  frames  and 

leaves.  The  filter  cloths  are  pierced 
with  openings  at  the  channels  for 
the  circulation  of  the  pulp  and  the 
filtrate.  These  cloths  are  large 
enough  to  extend  beyond  the 
frames.  The  follower,  a  solid  plate 
at  the  right,  is  now  brought  against 
the  assembled  parts  and  the  joints 
tightly  compressed  by  the  follower 
screw. 

Filling  is  generally  done  by  a 
three-throw  pump,  which  will 
charge  a  press,  forming  a  2-in.  cake 
in  from  eight  to  fifteen  minutes 
according  to  the  thickness  of  the  pulp  from  the  last  agitating  vat.  The 
final  pressure  is  as  much  as  60  Ib.  per  square  inch.  In  operation,  the 
pulp  flowing  along  the  left-hand  channels  of  the  frame  finding  no  out- 
let, the  solution  passes 
through  the  filter  cloths, 
leaving  the  solids  behind. 
The  solution  from  each 
leaf  flows  out  through  a 
cock  at  the  left-hand 
lower  corner  into  a  laun- 
der set  beneath  and  thence  TC 
to  the  gold  solution  tanks. 
Washing  the  cakes  is  the 
next  operation,  taking 
thirty  minutes  at  75  Ib. 

pressure  per  square  inch.  _ 

The  wash-solution  goes  to  j^^v&^/r 

the    wash-solution    vats. 

_.  ,  .    ,  FIG.  116A. — Merrill  Press-plate. 

The  cakes  are  now  dried 

by  compressed  air  for  about  two  minutes,  the  press  is  opened  and  dis- 
charged. A  cycle  of  operations  takes  seventy-two  minutes. 

The  Merrill  press  was  the  outcome  of  treating  low-grade  slime,  and  at 
the  Homestake  this  is  filled  into  the  presses,  the  gold  dissolved  in  them  and 
the  residue  automatically  discharged  without  the  apparatus  being  opened, 
except  for  necessary  repairs.  It  has  been  very  successful  at  this  mine  and 


PRESSURE  FILTRATION  175 

others  in  North  America.  In  principle  the  Merrill  press  is  similar  in 
many  respects  to  the  Dehne.  The  Merrill  is  essentially  of  the  ordinary 
rectangular  flush-plate-and-distance-frame  pattern  with  internal  channels, 
but  equipped  with  the  automatic  discharging  device  which  is  the  dis- 
tinguishing feature  of  the  press.  A  standard  press  to  hold  25  tons  of  slime 
contains  up  to  ninety-two  frames  4  in.  thick,  each  with  a  cross-sectional 
area  of  25  sq.  ft.  Between  the  frames  are  the  usual  solution  plates,  and  all 
are  made  with  suitable  channels  for  flow  of  solutions.  When  empty,  a  press 
weighs  about  70  tons.  The  filling  and  washing  is  similar  to  the  Dehne, 
only  pressures  are  about  half  of  that  used  in  the  latter  type.  Slime  may  be 
leached  hi  the  press  or  not,  according  to  the  character  and  value  of  the  ore. 
In  discharging  a  Merrill  press  the  procedure  is  as  follows:  In  a  lower  central 
channel  in  the  frames  is  a  3-in.  pipe,  resting  on  supports,  bolted  to  plates 
at  intervals,  throughout  the  length  of  the  press  and  connecting  at  the 
front  standard  with  the  water  supply  and  rotating  mechanism  for  the 
pipe.  Projecting  into  each  frame  compartment  from  the  sluicing  pipe  is 
a  0.16  in.  nozzle  through  which  water  is  discharged  against  the  cake  of 
slime,  while  the  whole  pipe  is  rotated  through  an  arc  of  about  200°.  The 
cycle  of  operations  at  the  Homestake  mill  is  560  minutes  and  at  Santa 
.Gertrudis  mill  ninety  minutes;  at  the  former  the  slime  is  leached  in  the 
press  and  at  the  latter  it  is  previously  agitated. 

As  typical  of  the  operating  cost  of  a  large  Merrill  filter  installation  in 
Mexico,  may  be  taken  the  following  figures  from  the  Esperanza  Mining 
Company  at  El  Oro.  Approximately  1000  tons  of  slime  are  filtered  daily 
with  six  presses,  averaging  eighty-two  frames  each,  being  equivalent  to 
100  Ib.  per  square  foot  per  day.  Caking  effluent  carries  $3.20  in  gold  and 
1  oz.  silver.  Dissolved  metal  loss  is  $0.03  in  gold  and  0.01  oz.  silver. 
Operating  charges  are:  Canvas,  $0.0108;  acid,  0.0035;  labor,  0.02;  mis- 
cellaneous, 0.00005;  sluicing  water,  0.0102;  a  total  of  $0.045  per  ton  fil- 
tered. The  item  for  sluicing  water  represents  all  charges  incidental  to 
settling  and  returning  for  re-use  all  water  sent  out  with  sluiced  residues. 

As  illustrating  the  additional  dissolution  of  metal  which  almost  invari- 
ably occurs  during  washing  in  pressure  filters  of  this  type,  the  following 
data  from  the  Merrill  installation  at  the  mill  of  the  San  Luis  Mining  Com- 
pany, in  Durango,  may  be  cited :  The  caking  effluent  carries  25  oz.  silver  per 
ton,  while  re-washed  filter  heads  and  tails  show  4.7  and  4.08  oz.,  respectively. 

GENERAL  REMARKS  ON  FILTERS 

For  large  output  are  used  the  various  types  of  suction  filters,  whether 
intermittent  like  the  Butters,  or  continuous  like  the  American  or  the  Oliver. 
Where,  however,  the  temperature  of  the  solution  approaches  the  boiling- 
point  they  are  not  effective,  due  to  the  vapor  generated  under  vacuum. 


176  CYANIDING  OF  GOLD  ORES 

Their  highest  suction  is  not  to  exceed  12  Ib.  per  sq.  in.,  but  this  may 
be  an  advantage  due  to  the  fact  that  the  lower  the  suction  the  more  open 
the  porous  structure.  Colloidal  slime  is  very  apt  to  pack,  making  slow 
filtering.  One  may  note  that  in  the  American  suction  filter  the  mud  layer 
is  automatically  washed  in  the  revolution  and  is  scraped  off  against  an 
inflated  filter  cloth  at  the  critical  instant. 

The  filter  presses,  such  as  the  Merrill  or  the  Kelly,  work  under  a  high 
pressure  of  40  to  50  Ib.  to  the  square  inch,  and  so  act  rapidly,  especially 
on  granular  material.  They  work  well  on  hot  solutions.  In  the  case  of 
the  Kelly  press  the  frames  must  be  washed,  then  withdrawn  and  unloaded. 
Sometimes  the  press-men  get  careless  and  do  not  wash  the  cake;  this 
results  in  loss  of  valuable  solution.  With  less  slime  proportionately  to 
handle  they  are  at  their  best,  since  stoppages  to  unload  are  less  frequent. 
The  Merrill  press  has  the  advantage  that  it  may  be  unloaded  ready  for  the 
formation  of  a  fresh  cake  without  having  to  open  up.  The  press  is  well 
adapted  to  clarification  or  to  filtering  precipitate  from  cyanide  solution, 
since  the  amount  filtered  out  is  small  compared  with  the  total  bulk  of  the 
solution.  Washing  and  air  drying  can  be  well  done  in  the  press. 

CLARIFYING 

The  solution  from  decantation  or  from  filter  pressing  is  not  always 
quite  clear.  A  little  slime  may  get  into  it  by  leakage  through  the  filter 
cloth,  or  an  overflow  may  at  times  be  turbid.  Such  solution  may  be 
passed  through  a  sand  filter,  vacuum  filter,  or  a  filter  press  of  a  few  leaves ; 
with  little  solid  matter  this  is  quite  readily  done,  and  leaves  a  bright 
and  sparkling  liquid.  If  the  solution  is  not  quite  clear  a  little  slime 
may  deposit  on  the  precipitant,  whether  zinc  shaving  or  zinc  dust,  thus 
diminishing  the  activity  of  precipitation. 

Of  the  filter  presses  the  Merrill  occupies  a  small  floor  space,  has  a 
capacity  of  1000  tons  of  solution  daily  and  is  readily  cleaned  without 
opening  the  frames.  To  this  end  the  press  is  sluiced  out  every  six  hours, 
using  barren  solution.  The  filter  cloths  are  of  No.  10  duck,  and  last  six 
weeks.  They  get  coated  with  a  lime  and  alumina  deposit,  so  that  .they 
must  receive  a  HC1  acid  treatment  about  every  three  days.  To  do*  this 
an  0.9  per  cent  acid  solution  is  pumped  through  the  press  at  a  pressure  of 
50  Ib.  for  an  hour.  This  is  withdrawn,  and  stored  for  re-use,  while  the 
press  receives  a  water  wash.  The  slime  caught  by  the  press  is  sent  back 
to  one  of  the  thickeners,  since  it  still  has  gold  values  and  cyanide. 

Where  sand  filters  are  in  use  an  extra  tank  is  required  in  reserve.  A 
sand-filter  plant  of  ample  capacity  occupies  a  large  floor  space  and  is 
expensive  to  erect. 

Unquestionably,  the  handling  of  slime,  as  produced  from  the  operation 


CROWE  VACUUM  PROCESS  177 

of  the  cyanide  process,  has  reached  a  point  where  no  further  radical  changes 
should  be  expected.  The  operation  of  separating  the  dissolved  metal  from 
the  treated  residue  may  now  be  carried  out  at  a  cost  of  $0.05  per  ton,  under 
favorable  conditions,  and  recoveries  of  98  and  99  per  cent  are  not  at  all 
unusual.  The  mechanics  of  the  various  machines  will,  of  course,  be  modified 
and  improved  to  suit  the  needs  of  special  problems,  and  there  is  always  the 
possibility  of  reducing  capital  cost.  In  comparison,  however,  with  the 
development  of  the  last  ten  years,  these  further  modifications  will  be  rela- 
tively unimportant. 

THE  CROWE  VACUUM  PROCESS 

This  consists  in  removing  from  the  solution,  just  prior  to  precipitation, 
substantially  all  the  dissolved  air  and  oxygen.  Generally,  cyanide  solu- 
tion going  to  precipitation  is  saturated  with  air  absorbed  during  the  air 
agitation  of  the  slime  pulp,  this  air  containing  30  per  cent  of  oxygen  and 
65  per  cent  of  nitrogen  as  against  23  per  cent  of  oxygen  and  75  per  cent 
of  nitrogen  as  found  in  ordinary  air.  Indeed  in  zinc  boxes  exposed  to  cold 
in  winter  a  white  precipitate,  a  hydrated  zinc  oxide,  will  form,  due,  doubt- 
less, to  the  dissolved  oxygen  of  the  solution.  In  presence  of  the  zinc  pre- 
cipitant the  oxygen  polarizes  or  reverses  the  action  of  the  precipitating 
couples  and  causes  re-solution  of  the  precipitated  gold  (or  silver).  Com- 
plete precipitation  can  take  place  only  when  sufficient  hydrogen  has  been 
evolved  to  combine  with  the  free  oxygen,  and  in  this  evolution  results  the 
consumption  of  both  zinc  and  of  cyanide. 

The  Crowe  apparatus  consists  of  a  receiver  or  drum  4  ft.  diameter  by 
10  ft.  high.  The  solution,  coming  from  a  steady-head  tank  set  24  ft. 
above  the  top  of  the  receiver,  pours  through  the  top  downward  over 
a  series  of  perforated  trays.  This  breaks  it  up  into  a  spray.  A  vacuum 
pump  continually  sucks  away  the  occluded  air  of  the  solution.  The  solu- 
tion level  at  the  bottom  of  the  receiver  is  maintained  .30  in.  deep  by  a 
float  operating  a  butterfly  valve  in  the  intake  pipe.  The  solution  flows 
away  at  the  very  bottom  to  a  pump  set  33  ft.  below  so  that  one  may  be 
sure  that  the  pump  foot-valves  are  covered  and  that  no  air  can  be  drawn 
in  when  pumping.  It  is  into  the  suction  pipe  of  this  pump  that  the  zinc- 
dust  is  introduced  for  precipitating.  It  is  computed  that  by  using  this 
process,  a  saving  of  50  per  cent  of  the  zinc  dust  can  be  effected. 

THE  PRECIPITATION  OF  GOLD  FROM  CYANIDE  SOLUTIONS 

Precipitation. — We  have  our  choice  between  the  ordinary  extractor- 
box,  using  zinc-shaving,  and  the  Merrill  precipitation  press,  using  zinc-dust. 

Zinc-boxes  are  efficient,  but  occupy  considerable  floor-space,  and  are  not 
simple  when  it  comes  to  the  clean-up.  There  is  always  the  aggravating 
problem  of  the  "  zinc  shorts." 


178 


CYANIDING  OF  GOLD  -ORES 


Merrfll  presses  are  compact,  thief-  and  fireproof,  and  allow  a  quick 
clean-up.  In  precipitating  strong  (0.2  to  0.3  per  cent  KCN)  and  rich  solu- 
tion with  concentrate  treatment,  they  give  good  satisfaction. 


THE  MERRILL  PRECIPITATION  PROCESS 

This  consists  of  the  introduction  of  zinc-dust  into  the  suction  of  a  pump 
from  a  pregnant  solution  tank,  or  the  flow  from  the  Crowe  Vacuum  treat- 
ment, by  means  of  a  special  feeder  through  a  filter  press.  Fig.  117  shows 


FIG.  117. — Merrill  Precipitation  Apparatus. 

the  preferred  type  of  feeder  for  zinc  dust.  This  is  contained  in  a  hopper 
marked  "  precipitant  feeder."  By  means  of  a  screw  feed  the  dust  is  fed 
in  an  accurately  adjusted  manner  through  a  small  hopper  into  the  "  pre- 
cipitant mixing  cone  "  there  mixing  with  a  stream  of  solution.  The  mix- 


MERRILL  PRECIPITATION  PROCESS  179 

ture  is  maintained  at  the  proper  height  by  means  of  a  float-controlled  valve 
at  the  bottom  of  the  cone. 

The  design  and  construction  of  a  satisfactory  feeder  is  considerably 
more  difficult  than  might  appear  at  first  sight.  The  original  type  as  pro- 
posed by  Merrill  was  an  endless  belt.  On  the  top  surface  of  this  belt  was 
placed  the  amount  of  zinc  necessary  to  precipitate  the  given  tank  of  solu- 
tion. The  belt  was  actuated  by  means  of  a  series  of  floats,  so  as  to  cause  the 
belt  to  travel  forward  and  feed  the  zinc  dust  hi  exact  proportion  to  the  rate 
at  which  the  solution  was  lowered  in  the  sump  canks.  A  later  type,  illus- 
trated in  Fig.  117  makes  use  of  a  screw  located  at  the  bottom  of  a 
hopper,  the  hopper  being  provided  with  a  hammering  arrangement  or  a 
reciprocating  arm  to  prevent  arching  of  the  zinc  dust.  Provision  for 
regulating  the  speed  of  the  screw  is  made  by  means  of  double  cone  and  belt 
drive. 

Operation. — The  rate  of  zinc  feeding  is  checked  by  weighing  the  amount 
run  in  hi  five  minutes,  experience  deciding  how  much  will  be  needed  accord- 
ing to  the  assay  value  of  the  solution.  When  a  newly  cleaned  press  is 
"  cut  in,"  that  is,  brought  into  use,  20  Ib.  of  zinc-dust  is  added  at  once, 
and  for  six  hours  the  rate  of  feed  is  doubled,  all  to  form  a  zinc-dust 
coating  upon  the  filter  cloths  of  the  press.  Also  the  solution  for  the 
first  fifteen  minutes  is  returned  to  the  gold  or  pregnant  solution  tank  as 
having  been  imperfectly  precipitated.  Where  two  presses  are  operated 
the  flow  is  turned  to  the  first  one  until  it  is  normal,  then  the  flow  is  through 
both. 

The  Clean-up. — The  press  is  dressed  with  four  thicknesses  of  cotton 
sheeting,  the  outside  cloth  being  removed  at  each  cleaning  (every  five  or 
six  days)  and  a  new  one  added  on  the  bottom.  To  clean  a  press,  the  solu- 
tion is  cut  off,  the  press  drained,  then  blown  with  compressed  air  for 
sixty  to  ninety  minutes;  this  dries  the  precipitate  to  about  45  per  cent 
moisture.  The  press  is  opened  and  the  precipitate,  amounting  to  perhaps 
130  Ib.  is  scraped  into  the  precipitate  wagon  set  beneath.  The  outside 
cloths  are  burned,  the  resulting  ashes  being  added  to  the  rest. 

The  barren  solution  from  the  Merrill  presses  is  returned  to  No.  4  thick- 
ener. 

The  method  of  precipitation  employed,  assuming  equal  efficiency  as 
regards  the  precipitation  and  recovery  from  solution  of  the  precious  metals, 
may  exert  an  important  influence  upon  extraction,  either  through  intro- 
duction of  the  precipitant  used  into  solution  or  its  failure  to  precipitate 
certain  interfering  elements.  For  example,  zinc  in  the  presence  of  arsenic 
may  interfere  in  the  treatment  of  certain  ores.  If  aluminum  precipitation 
is  used  the  difficulty  is  overcome  through  the  elimination  of  zinc.  When 
copper  reaches  a  certain  concentration  in  solution  difficulties  arise  with 
both  extraction  and  precipitation.  Neither  zinc  nor  aluminium  precipitate 


180 


CYANIDING  OF  GOLD  ORES 


copper  to  any  extent,  hence  if  copper  is  to  be  removed  from  solution  elec- 
trolytic precipitation  must  be  used. 

Carbon,  which  at  times  occurs  in  gold  and  silver  ores,  may  occasion 
difficulty  in  cyanidation.  It  has  been  generally  assumed  that  carbon  occurs 
in  gold  and  silver  ores  in  the  form  of  graphite,  but  the  evidence  available 
by  no  means  supports  this  view  in  all  cases.  The  two  extremes  of  carbon 
as  regards  its  behavior  in  cyanide  solutions  are  graphite  and  charcoal. 
Graphite  is  dense  and  does  not  possess  pores,  therefore  cannot  occlude 
gases,  while  charcoal  is  porous  and  has  the  property  of  occluding  relatively 
large  volumes  of  gases.  Graphite  does  not  precipitate  gold  and  silver  from 
cyanide  solutions  while  charcoal  does.  Intermediate  between  these  two 
extremes  are  various  forms  of  carbon  which  will  precipitate  gold  and  silver 
to  a  greater  or  less  extent. 

Feldtmann  has  shown  that  the  graphitic  or  carbonaceous  schist  from 
the  mines  of  West  Africa  will  precipitate  gold  from  cyanide  solutions 
roughly  in  proportion  to  the  carbon,  content  of  the  schist.  Gold  so  pre- 
cipitated is  not  soluble  to  any  appreciable  extent  in  fresh  cyanide  solution, 
but  is  soluble  in  sodium  sulphide  solution.  He  thinks  that  the  compound 
formed  is  possibly  carbonyl  aurocyanide,  which  may  react  with  sodium 
sulphide  according  to  the  following  equation  : 


(12) 


However,  the  graphite  was  found  definitely  to  interfere  with  cyanida- 
tion. Experiments  with  deflocculation  of  the  colloidal  portion  of  the  ore 
including  the  graphite,  gave  negative  results,  but  it  was  discovered  that  if 
the  physical  state  of  the  graphite  were  altered,  most  conveniently  by 
heating,  the  extraction  was  wonderfully  improved.  The  results  below 
show  clearly  the  effect  of  heating  : 

TABLE   I 


Per  Cent 
of  Gold 
Extracted. 

Pounds  of 
Cyanide 
Consumed. 

Per  Cent 
of 
Graphite. 

Wet  slime 

36  14 

0  9 

t65 

Dried  at  100°  C  

64.60 

0.6 

Check      

64  60 

0  6 

Heated  to  300°  C  

80 

0  7 

1  55 

Check 

83  00 

0  5 

Wet  slime  boiled  with  water,  graphite  skimmed  off. 
Wet  slime  boiled  with  water,  graphite  not  skimmed 
off 

56.00 
54  60 

0.6 
1  32 

Roasted  in  open  dish      

90.10 

0  12 

Roasted  in  open  dish                                   

90.10 

0.24 

0  63 

PRECIPITATION  METHODS  181 

Attempts  have  been  often  made,  with  some  degree  of  success,  to  pre- 
cipitate the  gold  electrolytically  upon  plates  suspended  in  the  turbid  solu- 
tion. In  present-day  practice,  however,  the  solution  is  invariably  clarified 
before  precipitation  is  attempted.  The  solutions  from  sand-leaching 
are  generally  clear  enough  for  precipitation,  but  the  solutions  from  slime 
treatment  need  clarifying  after  filtration. 

When  the  slime  and  sand  treatments  are  combined  much  of  the  solution 
from  the  slime  treatment  can  be  clarified  and  built  up  in  value  by  using  it 
for  the  first  washes  upon  the  sand,  but  even  then  the  solutions  from  the 
sand  plant  should  be  later  clarified  before  attempting  precipitation.  This 
has  been  done  by  the  use  of  the  sand  filter,  but  this  is  of  limited  capacity, 
and  the  preferred  method  is  the  use  of  a  leaf  filter  of  the  necessary  size,  as 
already  described  under  head  of  clarifying. 

An  expedient  sometimes  used  has  been  to  fill  the  head  compartments 
of  the  zinc  box  with  excelsior  or  other  fibrous  material  to  act  as  a  filter. 
This  must  be  removed  at  frequent  intervals  and  washed.  However,  at 
present  the  Merrill  clarifying  press  is  preferred. 

Gold  dissolved  by  cyanide  solutions  is  recovered  by  passing  the  solutions 
through  zinc  shaving,  zinc  dust,  zinc  wafers,  aluminum  dust,  or  charcoal. 
The  last  precipitant  is  not  used  much,  as,  although  efficient,  it  is  rather  a 
nuisance  in  requiring  a  good  deal  of  attention  during  operation  and  at  clean- 
up. The  most  commonly  used  precipitant  is  zinc-shaving  arranged  in  a 
long  narrow  box,  in  which  the  solution  is  made  to  flow  up  through  the 
mass.  This  is  the  original  MacArthur-Forrest  process.  The  reactions 
involved  are  fairly  well  understood,  and  in  depositing  gold  the  equation  is 
as  follows: 

(13)         KAuCN2+KCN+Zn+H2O  =  K2ZnCN4+Au+H+KOH 

Generally  speaking,  gold-bearing  solutions  should  be  no  lower  than  0.03 
per  cent  KCN  strength  for  good  results,  and  if  lower,  a  drip  of  strong  solu- 
tion, or  lumps  of  KCN  added  at  the  head  of  the  boxes  will  strengthen  them. 
If  copper  has  been  dissolved  from  an  ore,  it  will  precipitate  on  the  zinc, 
thus  preventing  proper  deposition  of  gold  and  silver.  A  partial  remedy  for 
this  is  to  either  dip  the  shaving  in  a  10  per  cent  solution  of  lead  acetate,  or 
add  the  latter  solution  regularly  at  the  head  of  the  boxes.  In  fact,  a  little 
lead  acetate  is  at  all  times  quite  useful  in  the  boxes.  Also  the  addition  of 
strong  solution  will  delay  precipitation  of  copper.  Silver  ores  always 
make  more  precipitate  than  gold. 

THE  ZINC  OR  EXTRACTOR  BOX 

A  description  of  a  typical  apparatus  used  in  precipitation  is  as  follows : 
A  zinc  box  (Fig.  118)  contains  seven  compartments,  each  12X15X24  in. 
These  have  perforated  false-bottoms  of  sheet-iron  or  wire-cloth  that  hold 


182 


CYANIDING  OF  GOLD  ORES 


up  the  zinc-shaving  with  which  they  are  filled.  The  partitions  are  set 
alternately  up  and  down,  to  compel  an  upward  flow  of  the  gold-bearing 
solution  through  the  shaving,  and  to  bring  it  intimately  in  contact  with  and 
insure  the  precipitation  of  the  gold  upon  the  surface  of  the  zinc.  At  a 
in  the  side  elevation  the  solution  enters  the  box  through  a  pipe  m  at  the  left, 
passed  through  all  the  compartments,  flows  over  the  last  partition,  and  dis- 
charges through  a  down-turned  pipe  into  the  sump  tank.  The  box  is  set 
at  a  grade  of  J  in.  to  the  foot. 


ELEVATION  AND  SECTION 

FIG.  118. — Seven-compartment  Zinc-box. 


1, 

Supports  for 
Screen   --*'*' 

L 

* 

i 

< 

2X 

1 

1J^'  Eipe  with  Cap  or  Plug  Cock 

ID 

^  "Launder  of  >/'El. 

• 

FIG.  119. — Sections  of  a  well-designed  Zinc-box. 

In  Fig.  119  we  have  a  section  of  a  well-designed  zinc  box  and  launder  of 
sheet  steel  with  wooden  interior  partitions.  The  supports  of  three  screens 
(forming  the  false  bottom  through  which  the  solution  rises)  is  as  indicated. 
The  bottom  is  inclined  to  drain  to  an  outlet  which  itself  is  plugged  from  the 
inside. 

The  Clean-up  of  the  Zinc-boxes. — This  is  made  monthly  or  bi- 
monthly, according  to  the  bulk  of  the  precipitate  to  be  treated,  and  the 
need  of  realizing  values  for  operating  expenses. 

Cleaning  the  Zinc  Boxes. — At  the  time  of  the  clean-up,  only  one  zinc- 
box  is  taken  care  of  at  a  time,  the  flow  continuing  in  the  others.  The 


ZINC-BOX  CLEAN-UP  183 

flow  of  gold  solution  to  this  box  is  stopped,  and  water  is  run  in  to  displace 
the  solution  contained.  Beginning  in  the  first  compartment,  the  fine 
material  is  lifted  out,  while  the  shaving  is  agitated  in  the  water  with  the 
hands  protected  by  rubber  gloves.  This  is  not  done  roughly,  for  the  brittle 
shaving  would  be  unnecessarily  broken  and  the  water  would  be  black  with 
the  floating  precipitate.  The  plug  hi  the  side  (see  k  in  the  cross-section 
Fig.  1 19)  is  gradually  withdrawn,  and  the  accumulated  slime  and  water 
allowed  to  flow  into  the  launder  h.  The  plug  is  replaced  and  the  com- 
partment is  again  filled  with  water.  The  zinc  again  is  rinsed  and  rubbed, 
and  the  loosened  precipitate  once  more  drawn  off.  About  three  such 
washes  free  the  shaving  from  precipitate  and  short-zinc.  The  compart- 
ments are  thus  successively  cleaned  up,  and  the  shaving  from  each  compart- 
ment is  moved  toward  the  head,  and  in  the  last  compartment,  where 
needed,  replaced  by  fresh  shaving.  Finally  the  launder  is  cleaned  with 
a  hose,  and  everything  washed  into  an  acid-tank  or  clean-up  sump 
along  with  the  material  first  lifted  out  of  the  compartments,  and  sulphuric 
acid  added. 

When  all  the  boxes  have  been  cleaned,  the  precipitate  is  allowed  to 
settle  a  short  time  in  the  acid-vat  and  the  supernatant  liquid  is  siphoned 
into  a  settling  tank.  In  this  larger  vat  the  particles  of  precipitate  have 
opportunity  to  settle,  and  to  be  recovered  subsequently  in  the  filter-press. 
The  acid-vat  is  stirred  by  hand  with  a  wooden  hoe,  or  preferably  a 
power-driven  agitator  in  constant  motion.  This  insures  a  thorough 
agitation  of  the  sludge  and  precipitate  hi  the  acid  treatment. 

The  Acid  Treatment. — Upon  the  watery  slime  about  30  Ib.  of  sul- 
phuric acid  is  poured.  This  acts  upon  the  short  zinc  and  produces  a 
violent  effervescence.  After  subsidence  the  whole  is  stirred.  When 
the  action  again  abates  15  Ib.  of  acid  and  the  same  amount  of  hot  water 
are  added  with  occasional  stirring.  This  is  repeated  until  further  addition 
of  acid  produces  little  effervescence.  Then  the  mixture  is  allowed  to 
stand  two  hours,  and  a  portion  is  tested  with  more  acid  to  see  that 
decomposition  is  complete.  The  total  time  for  the  operation  is  four  to 
six  hours. 

Filter-pressing  the  Precipitate. — The  black  mixture,  containing  zinc 
sulphate  in  solution,  is  diluted  with  hot  water  to  within  a  few  inches  of 
the  top  of  the  vat.  The  whole  content  is  stirred  and  then  pumped 
through  a  lead-lined  filter-press,  see  Fig.  113.  The  vat  is  washed  and  the 
washings  are  also  pumped  through  the  press.  Finally  the  residue  in  the 
press  is  washed  with  hot  water  to  entirely  remove  the  zinc  sulphate. 

The  entire  precipitate,  having  been  transferred  to  the  press,  while  in 
this  position  is  washed  with  water  under  pressure.  The  water  is  followed 
by  compressed  air  to  dry  the  precipitate,  which,  after  this,  is  ready  to 
discharge.  To  discharge  the  press,  the  tightening-screw  is  slackened,  the 


184  CYANIDING  OF  GOLD  ORES 

follower  is  drawn  back,  and  the  frames  are  successively  separated.  The 
grayish  .black  residue  in  the  recesses  of  the  frame  containing  20  per 
cent  or  less  of  water,  drops  into  a  drying-pan  placed  beneath  the  press  to 
receive  it. 

Dressing  the  Boxes. — At  the  start,  compartments  Nos.  1  to  4  inclusive 
are  packed  with  fresh  zinc  shavings;  in  two  days  the  fifth  and  in  two  days 
more  the  sixth  compartment.  As  zinc  is  consumed  fresh  shavings  are 
added  and  this  dressing  of  the  boxes  is  more  frequent  as  the  time  of  the 
monthly  clean-up  approaches.  The  first  compartment,  however,  takes 
already-plated  long  filaments  from  the  next  compartment  and  on  top  of 
them  shavings  now  nearly  consumed,  the  so-called  short  zinc. 

DRYING  AND  REFINING  THE  GOLD  PRECIPITATE 

The  product,  still  damp,  is  transferred  to  pans  24X44  in.  by  4  in. 
deep.  A  pan  is  slid  into  a  cast-iron  muffle,  and  heated  until  the  pre- 
cipitate is  dry  and  finally  to  an  incipient  red.  It  is  then  removed, 
allowed  to  cool,  and  the  weighed  contents  cautiously  mixed  with  50  per 
cent  borax,  some  sand  and  soda.  Since  the  product  is  light  and  dusty, 
care  must  be  taken  in  handling  it,  and  for  fusing  it  must  be  put  carefully 
into  melting  furnaces.  The  molten  metal  is  stirred  in  the  crucible,  then 
poured  into  crucible  molds,  and  on  cooling,  the  slag  is  removed  and  the  gold 
remelted  into  an  ingot. 

REFINING  WITH  BICHROMATE 

Where  a  filter  press  is  not  used  a  good  method  is  to  proceed  as  follows: 
To  the  acid-treated  and  washed  precipitate  is  added  five  times  its  estimated 
weight  of  water,  also  sulphuric  acid  and  bichromate  of  potash,  but  each 
stirred  in  separately  in  the  proportion  of  four  parts  of  60°  acid  to 
one  part  of  the  solid  bichromate,  the  latter  first  dissolved  in  hot  water. 
Careful'  additions  are  made  until  a  slight  coloration,  still  showing,  tells 
that  enough  bichromate  has  been  added.  The  remaining  precipitate  is 
decanted,  well  washed,  and  dried.  It  is  transferred  to  a  clay  crucible 
with  an  addition  of  borax  for  fluxing  and  niter  for  oxidizing  carbonaceous 
matter,  is  melted  and  poured  into  an  ingot  free  from  zinc,  lead,  or  copper. 

Refining  in  a  Cupelling  Furnace. — As  practiced  at  the  Alaska- 
Treadwell  cyanide  plant,  the  acid-treated  precipitate,  containing  17  per 
cent  of  silver,  5  per  cent  lead,  14  per  cent  copper,  5  per  cent  zinc  and  20 
per  cent  insoluble  matter,  is  treated  on  the  hearth  or  test  of  an  English 
cupelling  furnace  (see  Fig.  278) .  The  furnace  is  charged  hourly  with  the 
precipitate  to  which  have  been  added  fluxes.  This  melts  down  into  the 
lead  bath  contained  on  the  hearth.  Here  it  is  oxidized,  yielding  slag  and 


REFINING  GOLD  PRECIPITATE  185 

lead-bullion,  which  are  tapped  off  intermittently  at  opposite  sides  of  the 
hearth  as  they  accumulate.  A  typical  charge  mixture  would  consist  of 
precipitate  100  lb.,  glass  22  lb.,  sodium  carbonate  25  lb.,  old  slag  80  lb., 
iron  turnings  15  lb.  Such  a  charge  would  yield  about  35  lb.  of  rich  lead, 
10  to  15  lb.  matte,  and  160  to  180  lb.  of  slag. 

Drying  and  Refining  at  the  United  Eastern  Cyanide  Plant. — After 
determining  the  moisture  content,  the  undried  raw  precipitate  is  fluxed 
with  11  per  cent  borax  glass,  11 J  per  cent  sodium  bicarbonate,  6  per  cent 
manganese  dioxide,  3.3  per  cent  ground  bottle  glass,  and  at  least  10  per 
cent  of  old  slag  shells  from  former  melts,  the  percentage  being  in  terms  of 
the  calculated  weight  of  the  dry  precipitate. 

A  precipitate  press  ordinarily  runs  from  five  to  six  days,  and  yields 
about  130  lb.  dry  precipitate.  In  resuming  precipitation  after  a  final 
clean-up,  one  press  is  given  the  entire  flow.  When  this  builds  up  a 
pressure  of  35  to  40  lb.,  the  second  press  is  opened  just  enough  to 
maintain  the  pressure  of  the  first  press  below  45  lb.  When  the  second 
press  reaches  20  lb.  pressure,  the  entire  flow  is  turned  into  it  and  the 
first  press  is  cleaned.  This  method  is  carried  on  until  the  end  of  the 
month,  when  a  final  clean-up  is  made.  The  solution  is  metered  by  a 
revolution  counter  on  the  triplex  pump,  which  is  calibrated  at  intervals 
with  a  known  tonnage  of  solution.  Solution  samples  are  taken  each  shift; 
the  heads  are  a  dip  sample  every  hour  and  the  tails  a  drip  sample  from  the 
barren  flow.  The  tonnage  and  the  average  solution  assays  give  the 
ounces  of  gold  precipitated  daily;  this  is  checked  monthly  against  the 
bullion  sold. 

Melting. — The  fluxed  wet  precipitate  is  placed  in  No.  5  paper  bags 
and  fed  to  an  oil-fired  No.  150  Case  tilting  furnace,  using  a  No.  100  long- 
lipped,  graphite  pot.  When  ready  for  pouring,  the  pot  contains  fifteen 
sacks  of  precipitate  and  yields  a  600-  or  700-oz.  button  and  some  40  to  50 
lb.  of  slag.  This  charge  is  poured  into  a  conical  mold  and  allowed  to 
set  a  few  minutes.  The  slag  is  then  tapped  through  a  hole  about  2  in. 
above  the  gold  button,  and  run  into  cold  water  for  granulation.  The  cold 
skull  of  shell  left  in  the  mold,  which  contains  most  of  the  shot,  is  put  back 
with  a  subsequent  charge.  The  granulated  slag,  which  carries  some  25  oz. 
of  gold  per  ton  and  as  much  silver  is  ground  in  a  small  ball  mill  and  con- 
centrated with  a  laboratory-sized  Diester  table.  The  resulting  slag  tails 
from  each  month's  run,  of  which  we  usually  have  less  than  400  lb. 
carries  a  total  value  of  about  $50.  This  is  shipped  to  the  smelter  once 
a  year. 

An  average  month's  run  will  show:  Crude  dry  precipitate,  21,000  oz. 
troy  or  1440  lb.  avoirdupois;  71  per  cent  bullion  yield,  14,936  oz.  troy  or 
1024  lb.  avoirdupois;  39.7  per  cent  gold  yield,  8538  oz.  troy  or  573  lb. 
avoirdupois;  21.4  per  cent  silver  yield,  4497  oz.  troy  or  308  lb.  avoirdupois. 


186  CYANIDING  OF  GOLD  ORES 

The  crude  precipitate  contains  from  6  per  cent  to  7  per  cent  of  zinc  and 
about  10  per  cent  of  lead,  the  latter  coming  from  a  lead  acetate  drip  added 
as  the  solution  leaves  the  clarifying  filter. 

The  bullion  buttons  are  remelted  and  cast  into  bars  weighing  about 
150  Ib.  each.  A  dip  sample  is  taken  with  a  10-gm.  clay  crucible  just 
before  pouring.  This  sample  is  granulated  in  cold  water  and  is  sent 
to  the  assayer.  The  bullion,  as  shipped,  has  an  approximate  fineness  of 
560  in  gold  and  301  in  silver. 

CAPITAL  COSTS  OF  SLIME  PLANTS 

It  may  be  of  value  to  cite  approximate  figures  from  Rand  practice, 
where  the  intermittent  decantation  and  pump  transfer  type  of  plant  has 
reached  such  high  development.  The  total  approximate  capital  costs  of 
slime  plants,  in  dollars  per  ton  of  slime  treated  per  twenty-four  hours  is  as 
follows:  150  to  225  tons,  $292;  300  to  450  tons,  $276;  600  to  900  tons, 
$260;  1200  to  1800  tons,  $243.  Such  slime  plants  provide  for  collecting 
and  two  washes,  two  days  being  available  for  collecting  and  treatment,  and 
include  complete  pulp,  solution,  decanting,  and  water  services,  together 
with  solution  clarifiers,  but  exclude  precipitation,  refining,  and  power 
plants.  It  must  be  noted  that  Rand  plants  require  very  little  expenditure 
for  buildings. 

COSTS  OF  DISSOLUTION  BY  SLIME  AGITATION 

Costs  of  dissolution  may  be  of  some  interest,  although  it  must  be 
remembered  that  such  costs  depend  on  numerous  widely  varying  condi- 
tions, such  as  character  and  grade  of  ore,  cost  of  labor,  power,  and  supplies, 
amount  of  extraction  effected  prior  to  agitation,  arrangement  and  capacity 
of  the  plant,  and  the  recovery  percentage  achieved.  Furthermore,  such 
accessory  steps  in  the  cyanide  process  as  fine  grinding,  filtration,  or  decan- 
tation, precipitation,  melting,  and  refining,  will  depend  on  the  dissolution 
method  in  use  and,  hence,  should  be  taken  into  account  in  comparing 
"  slime  agitation  "  with  "  sand  leaching  "  or  "  filter  treatment."  Then 
too,  the  higher  percentage  of  dissolution  obtained  by  "  slime  agitation  " 
must  be  weighed  against  the  value  of  the  dissolved  metals  and  cyanide  lost 
by  imperfect  filtration  or  decantation.  It  is  to  be  understood  that  the 
following  figures  are  merely  numerical  averages  of  compilations  from 
representative  plants  in  various  parts  of  the  world.  The  cost  of  dissolution 
per  ton  of  slime  is  made  up  of  the  charges  for  thickening,  agitating,  pro- 
portionate pumping,  supervision,  cyanide,  lime,  and  lead  salts. 


FILTRATION  VS.  DECANTATION 


187 


COST  OF  FILTRATION  OR  DECANTATION 


Dissolution. 

Precip.  and 
Refining. 

Total. 

Average  gold  ore 

$0  29 

$0  25 

$0  54 

Average  silver  ore  

0.83 

0  42 

1  25 

Included  in  the  above,  but  so  far  below  the  average  as  to  require  special 
notice,  is  the  intermittent  decantation  method  practiced  on  the  Rand,  where 
the  slime  is  particularly  amenable,  in  so  far  as  concerns  thickening,  disso- 
lution, and  chemical  consumption.  Figures  from  nine  Rand  slime-plants 
show  an  average  cost  of  dissolution,  decantation,  precipitation,  and  refining, 
of  24.5  cents  per  ton  of  slime,  the  extraction  averaging  85.9  per  cent  on 
slime  heads  of  $1.92.  Recent  Rand  slime  plants  have  adopted  filtration  in 
place  of  decantation,  and  obtain  extractions  ranging  from  92  to  95  per  cent. 
From  the  above  Rand  figures,  about  4  cents  should  be  deducted  for  precip- 
itation, refining  and  assaying. 


COMPARATIVE  AGITATOR  DATA 


M. — Bottom  Drive  Mechanical 
P. — Pachuca 
D.— Dorr 


Sp.  Grav.  of  Pulp  =  1.326 
Power  at  $5  per  H.P.  per  month. 
Labor  at  $3  per  eight-hour  shift. 


Type. 

Si»e,  Feet. 

Capacity, 
Dry  Tons 
of  Slime. 

Total 
Corrected 
Cost. 

Erected  Cost 
per  Ton  of 
Capacity. 

Power  per  100 
Tons  of  Slime. 

Total  Operating  and 
Repair  Cost  per  Ton 
of  Slime  per  Twenty- 
four  Hours,  500 

Ton  Plant. 

M 

30X12 

134 

$2395 

$17.87 

D—  3.0H.P. 

D—  1.8  cents 

D 

30X12 

134 

2510 

18.73 

M—  5.2H.P. 

P—  2.3  cents 

P 

15X45 

102 

1935 

18.97 

P—  6.9H.P. 

M  —  2.4  cents 

1 

.i 

J 

ADAPTABILITY  TO 

.2 

Agitator 
Requisites. 

Violence  of 
Agitation. 

Total  Operating 
Cost  per  Ton 

j|g 

Freedom  from 
Choking  or  8 
tling. 

Ease  of  Startini 
after  Shut-do 

"tl 

fij 

Continuous 
System  of 
Agitation. 

Selective 
Agitation. 

Operating 
Simplicity. 

Minimum  Loss 
Mill  Height. 

In 

P 

D 

M 

D 

P 

P 

D 

D 

P 

D 

order  of 

D 

P 

D 

P 

D 

D 

P 

P 

D 

M 

preference 

M 

M 

P 

M 

M 

M 

M 

M 

M 

P 

The  rich  lead  must  be  again  cupelled,  this  time  to  remove  impurities 
such  as  copper  and  lead,  using  an  air  blast  for  oxidizing.     A  fine  gold  bul- 


188  CYANIDING  OF  GOLD  ORES 

lion  is  produced  which  is  remelted  in  a  Faber  du  Faur  retort  furnace  (see 
Fig.  156),  to  yield  ingots  or  bars  of  880  fine  in  gold. 

At  the  refinery  of  the  Goldfield  Cons.  Co.  they  have  adopted  blast- 
furnace smelting  of  briquettes,  a  mixture  of  the  precipitates  100  parts, 
litharge  100  to  125  parts,  and  heavy  sulphides  concentrate  from  the  mill 
(S  0.35  per  cent;  SiO2  30  per  cent)  together  with  some  mill  sweepings. 
The  furnace  yields  a  lead  bullion  (cupelled  as  above  described,  also  copper 
matte  and  slag.  The  matte  and  slag  may  be  sold  to  a  regular  smelting 
works. 


CHAPTER  XVI 
TYPICAL  GOLD   MILL  PRACTICE 

The  following  pages  give  descriptions  of  various  cyanide  mills: 

CYANIDING  FREE-MILLING  POROUS  ORES    • 

These  include  ores  which  are  somewhat  porous  and  which  may  be 
cyanided  by  leaching  when  coarsely  crushed. 

THE  WASP  NO.  2  MILL,  SOUTH  DAKOTA 

The  ore  is  a  massive  iron-stained  quartz. 

Its  average  gold  content  is  $2.40  per  ton.  The  ore  is  mined  by  steam- 
shovel,  and  crushed  by  two  No.  6  and  one  No.  4  Gates  crushers,  to  IJ-in. 
size.  Four  sets  of  16X36-in.  rolls  reduce  the  ore  to  J-in.,  when  it  is  ele- 
vated to  storage  bins.  The  rock  is  fed  from  the  storage  bins  through  rack 
and  pinion  gates  to  an  18-in.  rubber  belt  conveyor,  mounted  on  a  frame 
which  moves  back  and  forth  on  an  18-in.  track,  so  that  it  can  dump  to  any 
one  of  the  six  individual  conveyors,  each  of  which  serves  one  vat.  A 
1  H.P.  motor  and  rope  drive  serves  the  main  conveyor,  and  a  40  H.P. 
motor  the  individual  conveyors  and  feeders.  There  are  six  leaching  vats 
12X32  ft.,  holding  420  tons  each,  fitted  with  the  usual  filter  bottoms. 
Cyanide  solution  is  pumped  to  the  vats  so  that  it  flows  under  the  filters 
and  up  through  the  ore.  Two  or  3  Ib.  of  lime  is  added  per  ton  of  ore  for 
protective  alkalinity.  The  solution  contains  5  Ib.  of  cyanide  per  ton. 
This  stands  for  twelve  hours,  is  drawn  off  and  followed  by  seven  weak- 
solution  washes.  Gold  is  precipitated  by  zinc  shaving.  Cyanide  con- 
sumption averages  about  0.4  Ib.  per  ton,  and  costs  are  67  cents  with  an 
extraction  of  76  per  cent. 

ORES   OF  CLAYEY  NATURE  BY  CYANIDING 

Any  ore  which  contains  a  high  percentage  of  alumina  is  difficult  to 
treat,  and  two  good  examples  of  this  ore  are  the  Buckhorn,  Nevada,  and 
Victorious,  Western  Australia. 


190 


TYPICAL  GOLD  MILL  PRACTICE 


Main  Shaft 

Grizzly 


tion 


Vat 


THE  VICTORIOUS  MILL,  WESTERN  AUSTRALIA 

The  lodes  of  the  Victorious  mine  are  of  soft  kaolinized  material,  through 
which  run  small  veins  of  ironstone  quartz,  which  carry  the  gold.  This  ore 
has  been  successfully  treated  by  the  following  process,  of  which  a  flow 

sheet  is  shown  (Fig.  120).     The  present  daily 
output  averages  320  tons. 

Coarse  Crushing. — The  ore  is  broken  by  a 
10Xl8-in.  Blake-type  rock-crusher  running 
250  R.P.M.  After  being  crushed  the  broken 
rock  passes  to  a  14-in.  Robins  belt  conveyor, 
set  at  an  angle  of  20°  and  is  distributed  into 
a  bin  by  means  of  a  Robins  14-in.  hand- 
tripper  with  a  double  chute.  The  total 
storage  capacity  is  600  long  tons.  The  bin 
is  fitted  with  rack  and  pinion  doors  and  steel 
chutes. 

The  ore  is  ground  by  four  5-ft.  Huntington 
mills  fed  by  four  ore-feeders,  driven  by  means 
of  a  short  belt  from  the  Huntington  mill  shaft. 
Cyanide  circulating  solution  is  used  for  crush- 
ing, and  mercury  is  used  in  the  mills  with  the 
present  ore.  The  product  of  the  mills  equals 
1  ton  of  ore  to  1  ton  of  solution,  and  is  dis- 
charged into  a  cement  launder  and  thence 
to  the  pump  well.  From  here  it  is  lifted  35  ft. 
to  the  top  of  four  sets  of  cone  separators  by  a 
duplex  plunger  pump.  The  sand  separated 
*s  conveyed  by  launders  to  a  pair  of  Wheeler 
FIG.  120.— Flow-sheet,  Victorious  Pans-  Here  more  coarse  gold  is  collected  and 
Mill,  Western  Australia.  the  sand  is  ground.  The  overflow  from  the 
grinding  pans  joins  the  main  body  of  pulp 

from  the  separators  in  a  collecting-box  situated  4  ft.  above  the 
thickeners,  from  which  it  is  distributed  to  four  pulp  thickeners.  These 
are  steel  vats  25  ft.  diameter  by  9  ft.  deep.  The  agitators  aret20  ft. 
diameter  and  6'  ft.  deep  and  the  arms  revolve  at  7  R.P.M.  From  the 
agitators  the  pulp  is  pumped  to  a  distributing  agitator  near  the  filters. 
Filtration  is  done  by  three  Ridgway  filters  whose  capacity  is  estimated  at 
300  tons  per  twenty-four  hours  on  this  ore.  This  machine  has  a  large  shaft 
carrying  two  arms  on  which  are  suspended  a  basket  of  filter-leaves.  This 
is  dipped  into  a  pulp  tank  to  form  a  cake,  then  turned  into  a  wash  tank,  and 
finally  discharged  into  a  hopper  in  the  center.  The  ore  averages  $7  per  ton, 
and  an  extraction  of  90  per  cent  is  obtained  at  a  total  cost  of  $1.10  per  ton. 


THE  CITY  DEEP  MILL 


191 


Roof  Lights 


THE  KOLAR  FIELD 

The  ore  mined  at  the  Kolar  field,  India,  is  free  milling.  The  per- 
centage of  pyritic  content  varies  on  the  different  mines,  but  taken  as  a 
whole  it  does  not  average  more  than  1.5  per  cent  and  for  this  reason 
the  sand  can  be  weathered  without  detriment.  The  weathering,  while 
oxidizing  the  pyrite,  and  freeing  the  gold  content,  is  further  advantageous 
in  freeing  the  sand  of  surplus  moisture,  reducing  it  from  12  to  16  per  cent 
down  to  about  3  per  cent.  This  converts  it  to  a  more  friable  state,  in 
which  any  slime  present  can  be  easily  powdered.  Having  no  water  to 
displace,  it  can  thus  be  easily  saturated  with  solution,  and  the  lixiviation 
and  subsequent  washings  more  thoroughly  and  quickly  carried  out. 

THE   CITY  DEEP  MILL 

This  treats  ore  from  the  reefs  on  the  Rand  which  may  be  described 
as  a  conglomerate  of  quartz  pebbles  cemented  together  by  a  matrix 
having  a  dark  bluish  appearance  when 
freshly  mined.  The  "  banket  "  carries 
up  to  75  per  cent  silica  and  2.5  per  cent 
pyrite.  The  ore  from  the  mine  is  divided 
into  two  classes  by  screening  through 
grizzlies;  the  fine  is  conveyed  by  a  20-in. 
belt  direct  to  the  main  ore  bin.  The 
coarse  is  taken  by  four  inclined  sorting 
belts,  see  Fig.  121,  where  the  waste  rock 
is  picked  out  by  hand,  each  belt  feed- 
ing three  crushers  with  12X24-in.  jaw- 
opening.  The  return  portion  of  each 
sorting  belt  receives  the  rejected  waste 
rock  and  delivers  it  to  a  belt  that  takes 
it  to  waste.  The  transport  of  ore  to 
the  mill  and  all  necessary  surface  work 
is  effected  by  heavy  electrical  locomo- 
tives using  2000-volt,  50-cycle,  3-phase  current.  The  mill  of  2200  tons 
daily  capacity  is  equipped  with  200  stamps  arranged  in  units  of  ten,  each 
unit  driven  by  a  50  H.P.  motor.  The  weight  of  the  stamps,  which  have 
long  heads  and  short  stems,  is  2000  Ib.  One  special  feature  is  that  there  is 
only  a  layer  of  J-in.  felt  between  the  mortar  bases  and  concrete  foundations. 
King-posts  are  entirely  dispensed  with,  the  concrete  foundations  being 
carried,  with  indented  steel-bar  reinforcing,  to  above  the  level  of  the  top  of 
the  mortar-box.  Each  cam-shaft  is  carried  by  a  steel  frame  and  rests  on 
eleven  bearings,  so  as  to  minimize  the  risk  of  breaking.  Stems  are  4  in. 
by  13  ft.  long  and  the  stamps  are  arranged  for  a  heavy  duty  if  necessary. 


Sorting  Belt 


Return  belt  for 
carrying  waste  rock 


192 


TYPICAL  GOLD  MILL  PRACTICE 


Each  battery  is  provided  with  four  Challenge  feeders.  Referring  to  dia- 
gram, Fig.  122,  the  watery  pulp,  after  elevation  by  a  sand  pump,  is  classi- 
fied in  Caldecott  diaphragm  cones,  Fig.  58,  and  the  underflow  delivered 
to  the  tube  mills.  There  are  nine  of  these,  5J  ft.  diameter  by  22  ft.  long 
driven  by  a  100  H.P.  motor.  The  tube  mill  discharge  is  conducted  to 
amalgamating  tables  in  the  gold-recovery  house  and  under  the  same  roof 
are  arranged  the  extractor  or  zinc  boxes,  the  clean-up  machinery,  strong 
room  and  the  refinery  so  that  all  operations  are  performed  in  one  building 
under  the  supervision  of  a  responsible  man.  The  pulp  from  the  table  is 
again  elevated  to  four  coarse  sand  classifiers,  the  coarse  underflow  of 
which  goes  back  to  the  diaphragm  cone  for  regrinding.  Meanwhile  the 
overflow  of  three  classifiers  passes  to  the  slime  separators,  the  over- 
flow to  slime  collectors  as  described,  the  sand  underflow  to  the  sand 
collectors  consisting  of  six  steel  vats,  50  ft.  diameter  by  10  ft.  deep, 
set  on  reinforced  concrete  supports.  A  24- in.  conveying  belt,  running 


Diaphragm  Cone 


Fine  Pulp  Heritor 


FIG.  122. — Grinding  and  Classifying  on  the  Rand. 

under  the  center  line  of  these  vats,  takes  the  sand  excavated  from  them 
by  a  Blaisdell  excavator  and  by  conveying  belts  removes  it  to  twelve 
leaching  vats.  The  sand  is  fed  to  these  leaching  vats  by  a  Blaisdell  dis- 
tributor, and,  after  leaching,  it  is  discharged  by  a  Blaisdell  excavator 
through  a  central  discharge  opening  at  the  bottom  of  the  vats  to  two 
24-in.  conveying  belts,  one  under  each  row.  A  cross  belt  delivers  it  to  a 
24-in.  inclined  belt  which  takes  it  to  the  tailings  dump,  where  it  discharges 
100  ft.  above  the  ground. 

The  slime  is  collected  in  four  conical-bottom  steel  vats  also  on  reinforced 
steel  supports.  From  there  it  is  taken  for  treatment  in  two  steel  co^ical- 
bottomed  air  agitator  vats,  32  ft.  diameter  by  38  ft.  deep  and  later  washed 
in  eight  steel  conical-bottomed  vats  70  ft.  diameter  by  16|  to  20|  ft.  deep. 
This  washing  is  done  by  decantation. 

It  is  thus  seen  that  from  start  to  finish  the  labor  of  the  South  African 
native  has  been  eliminated  as  far  as  possible.  The  various  products  either 
flow  from  point  to  point  by  gravity  or  the  fluids  are  pumped  and  the  solid 
products  are  transferred  by  mechanical  methods. 


CONSOLIDATED  LANGLAAGHTE  MILL  193 


THE  MILL  OF  THE    CONSOLIDATED   LANGLAAGHTE   CO., 
RAND,   SOUTH   AFRICA 

The  mill  is  designed  for  a  capacity  of  45,000  tons  per  month  of  twenty- 
six  days.  The  receiving  bin,  which  is  of  steel  and  concrete  construction, 
delivers  the  ore  on  to  three  30-in.  belts  running  at  a  speed  of  150  ft.  per 
minute.  Each  belt  discharges  over  a  short  grizzly  into  a  washing  trommel. 
The  undersize  from  the  grizzlies  and  trommels  discharges  on  to  a  30-in. 
belt,  which  delivers  on  to  the  main  30-in.  conveyor  belt  leading  into 
the  mill.  The  washings  from  the  trommels  and  the  sorting  belts  are  con- 
veyed by  launder  to  the  coarse  sand  pumps.  The  oversize  from  each  of  the 
trommels  is  delivered  to  a  36-in.  sorting  belt,  running  at  a  speed  of  40  ft.  per 
minute.  Waste  rock  and  tube-mill  pebbles  are  thrown  into  separate  bins 
under  the  sorting  belts.  Each  sorting  belt  discharges  to  jaw  crusher,  the 
product  from  which  is  carried  by  a  24-in.  belt  to  the  mill  conveyor  belt. 
The  dust  produced  in  breaking  the  ore  is  exhausted  by  a  fan,  and,  after 
spraying  with  water  is  delivered  to  the  launder  carrying  the  trommel  under- 
size. The  sorted  waste  is  transported  from  the  bins  by  trucks  to  the  dump. 
The  ore  is  elevated  at  an  angle  of  18°  by  a  30-in.  belt  conveyor,  running  at 
a  speed  of  300  ft.  per  minute,  and  is  distributed  over  the  3000-ton  mill  bin 
by  means  of  a  tripper.  The  tube-mill  pebbles  are  conveyed  in  a  similar 
manner  from  the  crusher  station  and  thence  to  the  pebble-storage  bin  at  the 
west  end  of  the  mill  building.  The  mill,  is  supplied  with  100  stamps  each  of 
1750  Ib.  Ten  stamps  are  driven  by  a  50-H.P.  motor  operating  through 
a  counter  shaft  to  two  cam-shafts,  each  of  which  carries  five  stamps. 

There  are  ten  6Xl6j-ft.  tube-mills,  each  driven  through  spur  gearing 
and  belt  by  a  100-H.P.  motor.  The  battery  pulp  from  each  ten  stamps 
gravitates  to  a  6-ft.  primary  cone  classifier  at  the  head  of  each  tube-mill, 
One  secondary  cone  of  smaller  size  is  provided  for  each  two  primary  cones. 
The  underflow  from  the  primary  cones  passes  to  the  tube-mills,  and  the 
overflow  passes  into  the  secondary  cones.  The  overflow  from  the  latter  is 
led  direct  to  the  fine-sand  pumps,  while  the  underflow  is  led  to  the  coarse- 
sand  pumps.  The  tube-mill  pebbles  are  brought  from  the  pebble-bin  by 
cars  and  are  deposited  into  a  hopper  placed  at  the  inlet  of  each  tube-mill. 
In  the  platehouse  there  are  forty  stationary  amalgamating  tables,  each 
5X7  ft.,  and  recovery  by  amalgamation  is  72  per  cent.  Each  tube-mill 
discharges  through  a  launder  on  to  a  set  of  four  tables.  Opening  out  of 
the  platehouse  is  the  clean-up  room,  which  is  equipped  with  two  retorts, 
two  bullion  furnaces,  three  amalgam  barrels,  batea,  and  amalgam  press. 
A  strong-room  is  also  provided.  The  pulp  from  the  plates  is  conveyed  by 
a  launder  to  a  reinforced  concrete  sump  having  a  capacity  equal  to  the 
product  of  ten  minutes'  crushing  in  the  mill.  The  pulp  is  elevated  by  10-in. 
pumps  with  12-in.  suction  and  is  distributed  to  the  tube-mill  primary 


194  TYPICAL  GOLD  MILL  PRACTICE 

cones  by  means  of  launders.  The  overflow  from  the  secondary  cones  passes 
as  mentioned  above,  direct  to  the  fine-sand  pumps.  These  pumps  are 
provided  with  suction  hoppers,  having  overflow  launders  leading  back  to 
the  coarse-pump  sump.  The  fine-sand  pumps  elevate  the  final  pulp  to 
the  sand  classifiers.  The  principle  of  double  classification  is  adopted,  the 
classifiers  being  cones  8  ft.  diameter  by  10  ft.  deep.  The  primary  cones, 
six  in  number,  are  fitted  with  rings  to  catch  candle  grease  and  wood  fiber 
from  the  mine.  The  underflow,  suitably  diluted,  passes  to  three  secondary 
cones,  which  are  fitted  with  water  regulators  giving  a  wet  underflow.  The 
underflow  from  the  secondary  cones  passes  to  three  sand-collecting  tanks, 
which  are  each  55  ft.  diameter  by  10  ft.  deep,  and  are  provided  with  periph- 
eral launders.  Butters  distributers  are  used.  The  overflow  from  the 
primary  classifiers  is  led  to  four  sand  return  cones,  each  7J  ft.  diameter 
by  5  ft.  10J  in.  deep.  The  percentages  of  sand  and  slime  are  39.5  and  60.5 


FIG.  123.— Steel  Tank,  70  ft.  diam.,  for  Rand  Practice. 

respectively.  The  overflow  from  these  cones,  together  with  the  overflow 
from  the  secondary  classifiers  and  the  sand  collectors,  goes  direct  to  the 
slime  collectors  and  the  underflow  is  returned  to  the  fine  sand-pumps  for 
reclassification. 

The  sand  is  discharged  from  the  collectors  by  bottom  discharge  doors 
on  to  belt  conveyors,  which  carry  it  to  a  Blaisdell  distributer,  placed  over 
the  treatment  vats.  There  are  twelve  treatment  vats,  each  40  by  9i  ft., 
and  the  treated  sand  is  discharged  into  cars  and  conveyed  to  the  dump  by 
mechanical  haulage.  Recovery  by  sand  treatment  is  12  per  cent,  i  The 
extractor  house  contains  zinc  boxes  and  is  provided  with  an  acid-treatment 
plant  and  smelting-room.  The  latter  contains  calcining  and  reverberatory 
furnaces,  ball-mill,  and  strong-room.  The  decantation  slime  plant  consists 
of  the  following  tanks:  two  collecting  tanks  70  ft.  diameter  by  12  ft.  with 
7-ft.  6|-in.  cone,  see  Fig.  123;  five  collecting  tanks  50  ft.  diameter  by  12  ft. 
with  6-ft.  cone;  two  treatment  vats  70  ft.  diameter  by  12  ft.  with 
6-ft.  cone;  one  intermediate  transfer  tank  50  ft.  diameter  by  12  ft.  with 
7-ft.  6j-in.  cone;  ten  treatment  vats  50  ft.  diameter  by  12  ft.  with 


THE  HOMESTAKE  MILL  195 

6-ft.  cone.  Recovery  by  slime  treatment  is  12  and  total  extraction  96  per 
cent.  The  overflow  water  from  the  collectors  is  led  to  a  return  water  sump, 
from  which  it  is  pumped  to  the  mill  supply  tanks  by  10-in.  pumps.  The 
residue  is  worth  23  cents  per  ton. 

THE  HOMESTAKE  MILL,  LEAD,  SOUTH  DAKOTA 

The  ore  is  a  garnetiferous  hornblende  schist  carrying  7  to  8  per  cent  of 
pyrite  and  pyrrhotite  together  and  of  the  value  of  $4  per  ton.  It  is  treated 
by  amalgamation  and  cyaniding.  The  coarsely  crushed  ore  is  fed  to  640 
stamps  having,  as  shown  in  the  flow  sheet,  Fig.  125,  an  output  of  nearly 
2800  tons  per  day.  Here  it  is  crushed  in  cyanide  solution  in  the  ratio  of 
eleven  parts  solution  to  one  of  ore.  A  tube-mill  is  here  introduced  to 
increase  the  fineness  of  the  stamp-battery  product.  The  battery  pulp 
passes  over  in  all  two-thirds  of  an  acre  of  amalgamating  plates  set  upon 
a  12J  per  cent  slope  of  1|  in.  per  foot.  With  inside  amalgamation  in  addi- 
tion to  these  plates  72  per  cent  of  the  gold  is  caught.  The  tailings  now 
flow  to  fourteen  gravity  cone  classifiers  /,  each  4  ft.  diameter  by  5  ft.  5  in., 
which  yield  an  overflow  going  to  the  sand  classification  system  and  an 
underflow  of  8  per  cent  of  the  ore  which  goes  to  the  regrinding  system  to  be 
reground  in  tube-mills.  All  the  pulp  is  of  such  a  nature  that  a  solution 
contact  of  some  four  to  eight  hours  is  ample.  Direct  filter  slime  treatment 
is,  perhaps,  the  most  efficient  and  satisfactory  practice  of  the  day.  The 
partially  thickened  slime,  in  water,  is  charged  directly  into  the  filters,  where 
the  slime  cake  is  in  an  excellent  condition  to  receive  preliminary  treatment, 
such  as  aeration,  solution  leaching,  and  washing.  The  entire  dissolution 
is  effected  in  the  filter,  with  a  minimum  amount  of  solution,  precipitation, 
consumption  of  chemicals,  power  and  labor,  and  loss  in  dissolved  gold. 
The  plant  is  very  simple  and  compact,  while  the  various  operations  are 
susceptible  of  accurate  technical  control.  The  pulp  from  the  grinding 
machines  unites  with  the  main  stream  of  pulp  before  arriving  at  the  second 
battery  of  classifying  cones.  Classification  of  crushed  products  into  sand 
and  slime  is  effected  by  means  of  four  series  of  sheet  iron  cones  with  50  to 
80°  slope  with  peripheral  overflow,  each  unit  discharging  at  the  apex 
through  a  short  cast  bushing.  Nearly  all  of  the  material  overflowing  from 
the  various  cones  of  the  classification  system  will  pass  a  200-mesh  screen, 
and  is  treated  as  slime.  It  is  thickened,  and  then  run  to  the  slime  plant 
storage  tanks,  lime  is  added  and  treatment  continued  as  described  later. 

The  Sand-plant. — The  prepared  sand  contains  40.5  per  cent  coarse 
particles  that  remain  on  a  100-mesh  screen ;  30.8  per  cent  middling,  between 
100  and  200  mesh,  and  28.7  per  cent  fine  passing  200-mesh.  This  leaches 
at  the  rate  of  3  or  4  in.  per  hour.  Before  the  sand  enters  the  leaching  vats 
it  receives  a  stream  of  milk-of-lime  which  has  been  prepared  by  being 


196 


TYPICAL  GOLD  MILL  PRACTICE 


stamped  in  a  one-stamp  battery  reserved  for  the  purpose.  From  4  to 
5  Ib.  of  the  lime  is  added  per  ton  of  sand.  The  classified  pulp  and 
lime  thus  mixed  pass  to  a  Butters  distributor,  which  can  be  transferred 
from  one  vat  to  another  by  an  overhead  trolley.  There  are  20  leaching 


vats,  each  44  ft.  diameter,  9  ft.  deep,  and  capable  of  holding  600  tons. 
The  vat  is  filled  with  water  and  the  sand  runs  in.  It  takes  nine  hours  to 
charge  the  vat,  and  treatment  lasts  five  days.  When  the  vat  is  filled  the 
ore  is  drained  and  a  series  of  washes  of  the  stronger  of  the  stock  solutions 


THE  HOMESTAKE  MILL  197 

(containing  0.14  per  cent  KCN)  is  run  in,  allowing  each  wash  to  drain 
off  below  the  top  of  the  ore  to  draw  in  air.  Besides  this,  air  is  introduced 
below  the  filter.  The  effluent,  its  strength  reduced  to  0.10  per  cent,  is 
run  to  the  two  weak-solution  precipitation  tanks  /,  /,  Fig.  125,  each  26  ft. 
diameter  by  19  ft.  deep.  After  this,  the  weak  solution  is  brought  upon 
the  charge  and  retained  two  days  more.  The  solution  escaping  during 
this  period  is  run  to  the  two  strong-solution  collecting  tanks,  e,e.  This  is 
followed  by  a  water  wash  which  finally  reduces  the  unextracted  gold  to 
5  to  7  cents  per  ton. 

The  charge  is  now  ready  for  sluicing  out.  This  is  done  by  two  men  in 
2J  to  3J  hours,  four  side  gates  and  one  bottom  gate  being  used  for  the 
purpose.  The  8-oz.  duck  filter-cloth  underlaid  with  another  of  cocoa 
matting  is  washed  clean.  The  vat  is  then  filled  with  water  and  is  ready 
for  the  next  charging. 

The  Slime-plant. — The  slime  pulp,  amounting  to  1600  tons  daily,  has 
an  average  value  of  91  cents  per  ton.  It  contains  3  tons  of  water  to  1  ton 
of  solid,  and  is  carried  two  miles  by  a  12-in.  pipe  at  a  grade  of  1.5  per  cent 
to  the  slime-plant.  Here,  two  small  vats  are  provided  for  slaking  lime. 
The  content  is  drawn  to  a  screen-bottom  box  where  the  undissolved  lumps 
separate.  The  box  overflows  into  an  agitator  from  which  the  milk-of-lime 
continuously  runs  into  the  main  slime-stream  at  the  rate  of  5  Ib.  of  lime  per 
ton  of  dry  slime.  Two  storage  vats,  26  ft.  diameter  and  24  ft.  deep, 
having  conical  bottoms  with  47°  sides,  receive  the  stream.  From  the 
bottom  of  these  storage  vats  the  slime-pulp  is  drawn  continuously 
through  a  10-in.  pipe  to  large  Merrill  filter  presses,  65  ft.  below,  to 
obtain  a  pressure  of  30  Ib.  per  square  inch.  The  11 -in.  main  extends 
the  whole  length  of  the  press-building.  Between  each  pah*  of  presses  the 
main  branches  into  10-in.  pipes,  which  in  turn  send  two  4-in.  branches  to 
each  press.  The  smaller  branches  connect  to  a  4-in.  passage  or  channel 
that  extends  along  the  center  of  the  top  of  the  filter-press  frames.  From 
the  channel  the  slime-pulp  flows  into  the  press.  There  are  ninety-two 
frames  each  4  by  6  ft.  and  4-in.  distance-frames  to  form  slime-cakes  4  in. 
thick. 

Collecting  is  by  revolving  distributors,  while  the  residues  are  sluiced 
out  with  1  ton  of  water  per  ton  of  sand. 

Total  cycle  approximates  seven  days.  Screen  sizing  of  sand :  35-mesh 
(0.0164  in.),  23  per  cent;  35-35  (0.0082  in.),  29  per  cent;  65-100,  21  per 
cent;  100-200,  18  per  cent;  200,  9  per  cent. 

Extraction  on  $3.15  sand  heads,  79.9  per  cent. 

Cost  of  sand  treatment  per  ton  of  sand,  exclusive  of  neutralizing,  pre- 
cipitating, refining  and  assaying,  23.5  cents. 


198 


TYPICAL  GOLD  MILL  PRACTICE 


COST  OF  SAND  TREATMENT  PER  TON  OF  SAND 


Rand 

Cents. 

Labor 6.54 

Assaying,  sampling 0. 32 

Lime 0 . 56 

Transferring,  discharging 8 . 62 

Cyanide 7 . 23 

Precipitation 1 . 56 

Zinc  shavings 1 . 28 

Clean-up  smelting 2 . 92 

Sodium  bisulphate 0 . 94 

Miscellaneous  supplies 2. 30 

Water 1.60 

Power  and  lights 6.48 

Miscellaneous 0 . 59 


Homestake 

Cents. 

Superintendence. 1 . 06 

Assaying 48 

Neutralization 1 . 54 

Transportation  . 0. 17 

Classification 1 . 18 

Treatment 8 . 59 

Precipitation  and  pumping  solutions  1 . 18 

Refining • 38 

Heating 0.51 

Miscellaneous 0.81 

Repairs 1 . 88 


Total..  ..17.78 


Total..  ..40.94 


TOTAL  COST  OF  HOMESTAKE  DIRECT-FILTER  SLIME  TREATMENT   PER 

TON  OF  SLIME 


Cents. 

Superintendence 0 . 93 

Assaying. 31 

Neutralization 2 . 03 

Thickening 34 

Treatment 10. 73 

Precipitating  and  pumping  solution.  1 . 65 

Refining 66 

Heating 17 

Miscellaneous.  .  .88 


Or,  combined  as: 


Works  labor 

Shop  labor 

Power 

Chemicals 6 . 78 

Miscellaneous. .  .   2 . 22 


Cents. 

6.70 

.84 

1.84 


Total . 


18 . 38 


Total..  ..18.38 


CONSUMPTION  PER  TON  OF  SLIME 

Lbs. 

Sodium  cyanide 0  161 

Zinc  dust 0. 118 

Lime 3.840 

HC1 .  .0.393 


Power  in  Kw.  Hrs. 


1.150 


Actual  cost  of  dissolution,  deducting  precipitation,  refining,  assaying,  and  one- 
third  of  superintendence  and  miscellaneous  items 0. 1516 

Capital  Cost. — Assuming  an  average  total  cycle  of  six  hours,  the  com- 
plete capital  cost  of  this  plant  including  classification,  thickening,  filters, 
precipitation,  refining,  pumping,  heating — all  enclosed  in  buildings — will 
approximate  $265  per  ton  of  daily  slime  capacity. 


THE  LIBERTY  BELL  MILL  199 


THE  LIBERTY  BELL  MILL,  TELLURIDE,  COLO. 

The  mine  produces  a  hard  ore  containing  85  per  cent  silica,  10  per  cent 
lime  and  about  4  per  cent  pyrite.  The  treatment  consists  in  amalgama- 
tion, concentration  and  cyanide  treatment  of  the  concentrates. 

The  ore  coarse-crushed  at  the  mine  to  3-in.  size  is  delivered  by  tramway 
into  the  battery  ore  bins,  see  the  plan,  Fig.  126.     The  mill  contains  eighty 
stamps  of  850  Ib.  each  in  eight  batteries  set  on  concrete  foundations.     The 
ore,  fed  from  the  bins  by  Challenge  feeders,  is  wet  stamped  to  10-  to  12- 
mesh  in  a  solution  containing  2  Ib.  cyanide  per  ton.     For  each  battery 
there  are  two  plates  4  by  8  ft.  set  across  the  flow,  the  first  plate  with  a  grade 
of  2J  in.  to  the  foot  the  second  If  sq.  in.     The  slower  flow  on  the  flatter 
plate  ensures  recovery  of  the  finer  gold  particles.     The  pulp  is  now  treated 
in  four  Richards  three-spigot  vortex  classifiers  followed  by  six  6-ft.  settling 
cones,  the  spigot  discharge  being  concentrated  on  18  Wilfley  tables.     The 
concentrate  is  reserved  for  special  treatment  by  fine  grinding  and  cyaniding 
while  the  tailings  pass  on  to  be  reground  in  three  tube-mills  5  ft.  by  22  ft. 
long.     The  discharge  of  the  tube-mills  is  re-amalgamated  by  being  flowed 
over  eight  sets  of    transversely  set  amalgamating  plates  where  the  re- 
maining gold  particles  are  caught,  and  reconcentrated  on  ten  Deister  slime 
tables.     The  concentrate  is  united  to  the  first  above  mentioned  and  the 
slime  tailings  are  combined  with  the  overflow  of  the  cones  into  one  pulp 
to  go  to  the  nine  33  ft.  diameter  by  10  ft.  deep  Dorr  thickeners,  where  ade- 
quate settling  of  the  slime  can  be  effected.     The  thickened  underflow  or 
spigot  discharge  of  these  goes  through  six  Hendryx  agitators,  17  ft.  diame- 
ter, operated  continuously  in  series,  the  discharge  of  the  last  one  being 
pumped  to  the  equalizer  tank  of  the  filter  system.     Here  it  is  drawn  off  as 
needed  into  any  one  of  the  five  Moore  suction  filters  resembling  the  Butters. 
The  flow  or  strong  solution  from  the  Moore  filters  passes  to  solution 
storage  thence  to  the  sixth  filter  used  as  a  clarifying  filter  and  so  on   to 
the  precipitating  house  where  the  zinc  boxes  are.     The  weaker  solution 
resulting  from  the  washing  of  the  filters  also  goes  to  the  clarifying  filter 
and  to  other  zinc  boxes,  the  discharge  being  pumped  to  a  tank  to  be  used 
in  the  Moore  filter  for  blowing  off  the  cake.    The  precipitate  from  the  zinc 
boxes  is  acid  treated  in  the  precipitation  house  and  with  the  retorted 
amalgam  of  the  plates  is  melted  in  the  melt  house  and  shipped  to  the 
Denver  mint. 

TREATMENT  OF  TELLURIDE  ORES 

This  class  of  ore  is  mined  at  Cripple  Creek  and  Kalgoorlie,  but  the 
percentage  of  tellurium  is  now  quite  small,  the  bonanzas  having  been 
worked  out  for  several  years. 


200 


TYPICAL  GOLD  MILL  PRACTICE 


Precipitation  House 

i 

I 

0 

£ 
"3 

a 

t      1 

THE  GOLDEN  CYCLE  MILL 


201 


THE  GOLDEN  CYCLE  MILL,  COLORADO  SPRINGS,  COLO. 

This  is  a  custom  mill  treating  Cripple  Creek  telluride  ores,  containing 
no  other  deleterious  metals,  by  roasting  and  cyaniding.  The  ores, 
varying  in  size  from  1.5  in.  diameter  to  fine  sand,  after  sampling,  are 
taken  by  belt  conveyors  to  one  of  the  three  large  bins,  called  bed- 


LEGEND 

Dry  Material 
Pulp 


— . Solutions 

— -Wash  Solution* 


-] 


Tails  to  Slime  Dam 

FIG.  127. — Flow-sheet  of  Golden  Cycle  Mill. 


ding  floors.  They  are  classified  into  the  grades  A  and  B,  according 
to  the  lime  content,  and  separately  and  uniformly  distributed  in  large 
beds  of  5000  tons,  so  as  to  ensure  a  good  mixture.  Class  A  contains 
SiC>2,  76.7  per  cent  and  CaO  1.57  per  cent;  class  B  all  over  that,  a 
typical  analysis  of  the  latter  being,  insoluble  matter  75.9  per  cent;  A^Oa 
3.4  per  cent;  Fe  3.5  per  cent;  CaO  5.1  per  cent;  S  1.8  per  cent;  MgO  1.1  per 
cent.  Class  A  resembles  it  except  in  the  lime  content.  From  the  bedding 
floors  the  ore  passes  by  belt  conveyor  to  six  ball-mills  and  is  then  dry 
crushed  to  pass  a  |-in.  screen  and  with  an  average  of  not  more  than  30  per 


202  TYPICAL  GOLD  MILL  PRACTICE 

cent  coarser  than  10-mesh.     The  basic  ores  need  finer  crushing  than  the 
silicious  ones.     The  united  capacity  of  these  mills  in  1250  tons  daily.     The 
ore  from  the  mills  is  carried  by  a  belt  conveyor  to  steel  bins  set  above  nine 
duplex  Edwards  roasters,  Fig.  74.     This  furnace  165  ft.  by  13  ft.  wide  has 
a  roasting  hearth  area  of  1495  sq.  ft.     The  rabbles  revolve  6  R.P.M. 
There  are  three  fireboxes  to  each  roaster.     There  is  a  cooling  hearth  44  ft. 
long  by  14  ft.  wide  or  with  an  area  of  572  sq.  ft.     In  roasting  class  A  ore, 
the  temperature  in  the  flow  of  heat  at  No.  2  firebox  is  800°  to  850°  C. 
For  the  class  B  ore,  a  higher  temperature  of  850°  to  900°  C.  is  maintained. 
The  ore  escapes  from  the  roasting  hearth  at  485°  C.  and  leaves  the  cooling 
hearth  at  278°  C.    For  roasting  class  A  ore  12.5  per  cent  of  fuel  is  used, 
class  B,  containing  so  much  more  lime,  takes  a  higher  heat  and  at  least 
17  per  cent  of  a  local  lignite  coal.     Each  furnace  roasts  125  to  150  tons  of 
class  A  ore,  80  to  100  tons  of  class  B  ore  daily.     The  roasted  ore  falls  from 
the  cooling  hearth  upon  a  reciprocating  drag  conveyor,  where  it  is  sprayed 
with  water  to  cool  it  to  about  90°  C.  (194°  F.),  so  that  it  may  fall  upon 
the  rubber  conveying  belt  to  be  taken  to  bins  set  above  seven  6-ft.  Chilian 
mills  where  the  cyanide  treatment  begins.     Here  the  ore  is  fed  with  addi- 
tion of  cyanide  solution  to  the  mills  and  ground  to  pass  a  50-mesh  screen. 
The  discharging  pulp  is  distributed  upon  the  tables,  blanket-covered  and 
set  at  a  slight  slope,  where  the  pulp  spreads  out  in  an  even  layer  and  where, 
with  the  aid  of  some  solution,  the  lighter  pulp  is  washed  away,  while  the 
coarse  gold  developed  in  the  roasting  is  caught  in  the  interstices  of  the 
blanket.     The  pulp  now  passes  to  the  bowl-type  Dorr  classifiers  where  it  is 
separated  into  two  products,  clean  sand  for  leaching,  and  slime,  which  is 
agitated  and  the  gold  solution  removed  by  vacuum  filters.     One  may  note 
here  the  beneficial  effect  of  roasting  for  not  only  is  the  telluride  of  gold 
decomposed,  but  the  colloids  of  flocculent  portions  are  shriveled  up  by 
the  heat.     Precipitation  is  effected  by  zinc  shaving  in  the  usual  manner. 
Cripple  Creek  ore,  if  slimed  and  given  a  long  period  of  agitation,  will 
yield  as  much  as  by  the  above-described  method.     Yet  roasting  prevails, 
since,  while  it  costs  50  to  60  cents  per  ton,  it  is  not  necessary  to  grind  so 
fine,  the  extraction  is  a  matter  of  hours  rather  than  of  days,  cyanide  con- 
sumption is  less,  and  the  solutions  are  less  likely  to  become  foul. 

| 

THE   VICTOR   PLANT    OF   THE   PORTLAND    GOLD    MINING    CO.,    VICTOR, 

COLO., 

was  built  for  the  treatment  of  the  ore  from  the  Portland  mine  which  would 
not  withstand  the  high  cost  of  freight  and  treatment  when  shipped  to 
the  Portland  plant  at  Colorado  Springs. 

The  ore  is  brought  to  the  mill  in  5-ton  electric  cars  and  dumped  into  a 
cylindrical  steel  bin  above  the  crushing  plant.  From  this  bin  it  is  fed  by 
an  apron  conveyor  to  a  15  by  30-in.  Blake  crusher,  which  reduces  the  ore 


THE  VICTOR  MILL  203 

to  about  3-in.  size.  It  then  passes  to  a  36-in.,  style  B  Symons  disk  crusher, 
which  machine  reduces  it  to  l£  in.;  thence  to  a  set  of  20  by  48-in.  rolls, 
the  entire  product  of  which  will  pass  a  1-in.  ring.  A  belt  conveyor 
takes  this  1-in.  product  to  the  main  mill  building,  where,  after  passing 
through  a  Vezin  sampler,  it  is  distributed  into  four  steel  storage  bins. 
These  four  bins  discharge  by  plunger  feeders  to  four  6-ft.  Akron  Chilean 
mills.  At  this  point  a  weak  cyanide  solution  is  introduced,  the  mills  dis- 
charging a  pulp  through  a  30-mesh  screen,  which  flows  and  is  distributed  to 
thirty-six  Wilfley  tables.  The  concentrate  from  these  tables  is  finished  on 
six  Wilfley  finishing  tables.  This  last  set  of  tables  yield  heads,  high  in 
iron  sulphide  and  containing  some  gold  which  goes  to  the  smelter,  and  a 
silicious  tailings,  which  after  sliming  in  a  tube-mill  is  mixed  with  the  regular 
mill  slime. 

The  tailing  from  the  Wilfleys  runs  to  four  Akins  classifiers,  where  it  is 
divided  into  sand  and  slime.  The  sand  goes  to  a  continuous  wash 
system  (Akin  classifiers)  whence,  after  being  washed  free  of  soluble 
gold,  it  is  hauled  to  the  dump.  The  slime  is  pumped  to  thickening  cones, 
where,  after  thickening  it  is  reconcentrated  on  Card  tables.  The  con- 
centrate from  the  Cards  join  that  from  the  Wilfleys.  The  tailing  from 
the  Cards  runs  to  the  Akins  thickeners,  the  thick  pulp  from  the 
same  going  to  air  agitators  and  thence  to  Portland  filters,  whence,  after 
being  washed  free  of  soluble  gold,  it  is  hauled  to  the  dump.  The  effluent 
solution  from  the  Portland  filters  joins  the  clear  overflow  from  the  thick- 
eners is  clarified,  treated  by  the  Crowe  vacuum  method  and  by  the  Mer- 
rill precipitating  process,  and  goes  to  the  zinc-dust  precipitating  plant. 
The  mill  has  a  capacity  of  500  tons  daily,  and  uses  only  1,000,000  gal. 
of  precipitated  water  per  month. 

KALGOORLIE  DISTRICT,  WESTERN  AUSTRALIA 

The  ores  are  silicious  and  contain  pyrite  and  tellurides  of  gold  and  silver. 
There  is  but  little  free  gold  and  that  mainly  in  the  pyrite.-  At  the  Kalgoor- 
lie  mills  two  systems  of  treatment  grew  up,  the  dry  or  roasting  process  as 
described  for  the  Portland  plant  at  Colorado  Springs  and  the  wet  process. 
It  soon  was  found  that  the  telluride  ores  were  not  soluble  in  ordinary 
cyanide  solutions,  but  bromo  cyanide  proved  to  be  effective  (see  "  The 
Bromo-cyanide  Process.") 

Grinding  pans  appear  to  hold  their  own  in  the  treatment  of  telluride 
ores  in  the  Kalgoorlie  district.  They  are  used  for  intermediate  grinding, 
but  for  sliming,  tube-mills  are  considered  preferable  to  pans.  The  For- 
wood-Down  pan  has  a  classifying  discharge,  the  pulp  issuing  from  the 
pan  by  a  row  of  1-in.  holes  near  the  top.  Outside  the  pan  is  a  pocket  or 
launder  having  a  slot  at  the  bottom  leading  back  to  the  pan.  Heavy  sand 


204  TYPICAL  GOLD  MILL  PRACTICE 

settles  in  this  pocket  and  is  returned  through  the  slot  while  the  fine  flows 
over  the  outer  edge  of  the  pocket. 

Gradually  the  filter-press  on  the  Kalgoorlie  field  is  being  discarded  in 
favor  of  vacuum-filters.  In  the  past  an  immense  sum  has  been  spent  on 
press  plants,  about  100  presses  being  erected,  75  of  which  are  treating  100,- 
000  tons  monthly.  The  others  are  out  of  commission,  but  the  benefit 
derived  from  this  machine  at  Kalgoorlie  has  been  admittedly  large.  With  a 
press,  washing  can  be  carried  to  a  degree  that  cannot  be  beaten,  but  the 
labor  cost  is  high,  and  it  is  expected  that  the  press  will  have  to  give  way. 
The  Associated  Northern,  Boulder,  and  Oroya  Links  filter  their  current 
mill  slime  by  vacuum  systems  of  their  own,  while  others  are  talking 
about  introducing  the  system. 

THE  OROYA-BROWNHILL  MILL,  KALGOORLIE  DISTRICT 

This  uses  a  wet  process  of  concentration  followed  by  cyaniding.  The 
ore  after  coarse  crushing  in  a  rock-breaker  goes  to  the  stamps.  Here  it  is 
crushed  in  a  weak  solution  of  about  0.4  per  cent  or  0.8  Ib.  per  ton  of 
cyanide,  kept  alkaline  by  the  addition  of  lime  and  the  pulp  run  to 
hydraulic  classifiers.  Here  in  closed  circuit  with  Wheeler  pans,  Fig.  135, 
the  sand  is  finely  ground,  the  classifier  overflow  then  going  to  Wilfley 
tables.  The  concentrates  from  the  tables  receive  a  special  treatment  as 
follows : 

These  amount  to  6  per  cent  of  the  ore  milled  and  contain  11  ounces 
gold  per  ton  and  the  great  bulk  of  the  refractory  elements  in  the  ore,  thus 
leaving  a  tailings  product  well  suited  to  subsequent  cyaniding.  The  con- 
centrate is  sent  to  three  single-deck  Merton  roasting  furnaces  similar  to 
the  Edwards  furnace  (Fig.  74),  where  it  is  roasted  with  from  0.5  to  2.2  Ib. 
salt  per  ton,  each  furnace  easily  roasting  10  tons  per  day.  The  roasted 
product  is  removed  by  push  conveyor  as  shown  in  the  Edwards  roaster 
to  two  pairs  of  Forwood-Downs  5-ft.  pans  in  parallel.  The  first  pair  is 
used  both  for  fine  grinding  and  amalgamation,  their  overflow  passing  to  the 
second  pair  used  for  fine  grinding  only.  Mercury  is  added  to  the  first  pans 
three  times  daily,  amalgamation  recovery  is  about  30  per  cent  of  the  value 
of  the  concentrate.  The  second  set  of  pans  give  an  overflow  whicji  passes 
on  for  cyanide  treatment.  It  is  agitated  in  vats  with  mechanical  agitators 
with  a  solution  of  0.1  per  cent  or  2  Ib.  per  ton  for  100  hours,  lime  and  lead 
acetate  being  sometimes  added  in  the  pans.  The  pulp  is  afterwards 
filter-pressed.  The  pregnant  solutions  are  precipitated  in  ordinary  zinc 
boxes  using  zinc  shavings,  that  have  been  dipped  in  lead  acetate  solution. 

Returning  now  to  the  tailings  from  the  Wilfley  tables,  these  are  classi- 
fied in  closed  circuit  with  tube-mills,  so  as  to  give  an  all-slimed  product  for 
cyaniding.  This  is  dewatered  or  thickened  and  goes  to  the  agitation 


THE  HOLLINGER  MILL  205 

vats  having  mechanical  stirrers.  Here  it  is  agitated  with  cyanide  solu- 
tion of  1  per  cent  or  2  Ib.  cyanide  per  ton  for  three  hours,  the  cyanogen 
bromide  is  added  at  the  rate  of  about  1  Ib.  for  each  ounce  of  gold  and  the 
agitation  continued  for  twelve  hours.  About  two  hours  before  the  com- 
pletion of  the  agitation  quick  lime,  2  or  3  Ib.  per  ton  is  put  in.  After  agi- 
tation the  pulp  is  discharged  to  a  stirrer  or  agitator  to  keep  it  property 
mixed  while  it  is  being  pumped  into  the  filter-presses.  The  gold-bearing 
solution  from  the  filter-presses  passes  through  clarifying  presses  and 
thence  to  the  zinc  boxes  for  precipitation.  All  the  tailings  are  taken  from 
the  presses  to  the  dump  by  conveying  belt.  In  1905  when  the  ore  carried 
$32  per  ton  in  gold  the  extraction  was  94.6  to  95  per  cent. 

THE  HOLLINGER   MILL,   PORCUPINE  DISTRICT,   ONTARIO,    CANADA 

This  is  one  of  the  new  mills  of  the  district  and  treats  499  tons  of  $20  ore 
daily,  recovering  93  per  cent  of  the  gold.  The  ore  is  of  quartz  and  schist 
with  a  high  percentage  of  pyrite.  It  is  soft  and  easily  crushed,  yielding  a 
heavy  pulp  which  gives  rise  to  some  mechanical  difficulties  in  the  agitation. 
The  gold  is  free.  Treatment  consists  in  concentration,  cyaniding  the 
concentrates  and  the  tailings.  Fig.  128  is  a  flow-sheet  of  this  mill.  The 
ore  is  coarsely  crushed  in  two  stages,  the  product  from  the  first  or  gyratory 
breaker  being  screened  by  a  trommel  having  2j-in.  holes.  The  oversize 
of  this  is  crushed  by  a  20-  by  10-in.  Blake  crusher  so  that  a  product  of  less 
than  2J  in.  passes  by  belt  conveyor  to  be  distributed  to  the  mill  bins. 
Thence  it  is  drawn  off  in  regulated  quantity  by  Challenge  feeders  to  the 
stamp.  It  will  be  noted  that  the  feed  bins  hold  2^  days'  supply  of  ore. 
Stamp  crushing  is  done  in  a  solution  of  1.5  Ib.  cyanide  per  ton,  using  5  tons 
per  ton  of  ore  crushed.  The  ore  is  now  crushed  to  6-mesh  size  through  Dorr 
classifiers  and  tube-mills  in  closed  circuit.  The  overflow  from  the  classifier 
is  first  again  treated  by  means  of  large  Spitzkasten,  20^  ft.  long,  6  ft.  Wide, 
and  6  ft.  deep  to  yield  two  products,  an  overflow  for  direct  cyaniding  and 
an  underflow,  a  product  that  undergoes  extensive  concentration  in  order  to 
recover  the  free  gold.  The  underflow  of  25  per  cent  solids  is  concentrated 
upon  Deister  slime  tables  similar  to  the  Wilfley  table  (Fig.  84)  yielding 
concentrates  largely  pyrite,  containing  the  free  gold  and  a  tailings  which 
joins  the  Spitzkasten  overflow  for  cyanide  treatment.  The  concentrate, 
comparatively  free  from  solution,  is  taken  by  a  spiral  screw  conveyor  and 
bucket  elevator  to  four  small  bins  ready  to  be  fed  into  Wheeler  pans  in 
charges  of  1J  tons  of  the  concentrate  with  100  Ibs.  of  mercury.  After 
grinding  for  several  hours  the  contents  of  the  pans  are  discharged  to  8-ft. 
settlers,  Fig.  136,  where  the  tailings  overflow  for  cyaniding  while  the 
amalgam  collects  in  the  mercury  well  of  the  settler.  Referring  now  to 
the  various  overflow  products  for  cyaniding,  these  are  pumped  to  Dorr 


206 


TYPICAL  GOLD  MILL  PRACTICE 


Ore  (370-400  Tens  per  Day) 


No.  7  Kennedy 


Gyratory 


Trommel  2,  5-In.  Holes 


20-InBeltConveyTr 

Croa3  Belt  Conveyor 

Mill  Bins  (1000  Tons) 


Battery  Storage 


1.5  -Ib.  KCN 

Solution 
6  Tone  person  Ore 


40-15W) 


8  Suspended 

Challenge  Feeders 


Ib.  Stan)]*,      narrow  Mortars 


FIG.  128.— Flow-sheet  of  Hollinger  Mill. 


THE  TOM  REED  MILL  207 

.j 

thickeners  which  deliver  an  underflow  containing  50  per  cent  solids. 
With  an  addition  of  barren  cyanide  solution  this  product  is  agitated  for 
forty-eight  hours  by  Trent  agitators  in  series  and  is  then  filtered  by  Moore 
suction  filters  to  give  a  tailings  that  is  rejected  still  holding  25  per  cent 
moisture.  The  clear  overflow  of  the  Dorr  thickener  is  joined  by  the  filtrate 
from  the  Moore  filters,  but  must  be  clarified  before  precipitation.  Pre- 
cipitation, using  zinc  dust,  is  performed  as  described  under  the  head  of 
"  Merrill  Precipitation  Process."  The  filtrate  returns  as  barren  solution 
to  the  battery  storage  tank  near  the  stamps. 

THE  TOM  REED  MILL,  OATMAN,  ARIZ. 

The  treatment  is  by  cyaniding  using  continuous  counter-current 
decantation.  The  gold  occurs  principally  as  hematite  in  quartz  ore 
of  low  grade,  but  hi  large  bodies.  Referring  to  the  flow-sheet,  Fig.  129,  the 
run  of  mine  ore  crushed  through  a  gyratory  and  a  Dodge  jaw-crusher  in 
series  is  reduced  to  a  maximum  size  of  2^  in.  and  is  taken  by  an  inclined 
belt  conveyor  to  the  feed  bins.  A  16  by  16-in.  steel  chute  at  the  bottom  of 
the  bin  makes  it  possible  to  draw  off  the  ore  to  a  Stephens-Adamson  apron 
feeder  to  a  6  ft.  by  5  ft.  ball-mill  which  grinds  in  closed  circuit  with 
a  Dorr  classifier,  with  addition  to  some  pregnant  cyanide  solution  from 
the  decantation  tanks.  The  overflow  from  the  classifier  passes  to  two 
pairs  of  similar  classifiers  on  the  floor  below,  each  pair  in  closed  circuit 
with  a  5  by  6-ft.  ball-mill.  Here  a  further  addition  of  pregnant  solution 
results  in  a  product  85  per  cent  of  which  is  less  than  200-mesh  size,  this 
fine  grinding  being  necessary  to  ensure  contact  of  the  gold  by  the  cyanide 
solution.  The  pulp  thickened  in  the  primary  40-ft.  Dorr  thickener 
to  a  specific  gravity  of  1.5  is  raised  to  the  first  of  four  40-ft.  agitators, 
where  it  is  agitated  in  series  with  addition  of  air.  From  the  last 
of  these  machines  the  pulp  passes  to  the  lower  of  two  Eronier  sand 
pumps.  The  lower  pump  delivers  to  the  upper  one  and  that  in 
turn  to  a  distributing  box.  From  this  box,  the  pulp  is  divided 
between  the  head  tanks  of  two  series  each  of  four  Dorr  thickeners, 
Fig.  103,  operating  on  the  continuous  counter-current  decantation  sys- 
tem (C.  C.  D.).  From  the  last  of  the  series  the  residual  tailings 
go  to  a  settling  pond.  The  overflow  from  the  head  tanks  flows  as  already 
mentioned  to  the  ball-mills  and  through  them  to  the  primary  thickener. 
The  nearly  clear  overflow  of  this  thickener  passes  on  to  a  vacuum  classifier 
filter  and  after  treatment  by  the  Crowe  vacuum  method  is  precipitated 
with  zinc  dust  by  the  Merrill  precipitating  process  in  a  building  near 
the  barren  solution  sump.  The  filtrate  from  the  Merrill  process  flows 
by  gravity  to  two  6  by  6-ft.  measuring  tanks  that  are  alternately  filled 
and  emptied  by  the  action  of  a  tilting  launder  operated  by  two  floats,  one 


208 


TYPICAL  GOLD  MILL  PRACTICE 


in  each  tank.  This  launder  in  turn  opens  the  discharge  of  a  full  tank  and 
closes  the  empty  one  while  at  the  same  time  it  diverts  the  flow  to  the  latter. 
The  number  of  tanksful  is  shown  by  an  automatic  counter.  From  these 


Allis-Chalmers.  Ball  Mills 

100  Hp..  6x5' 

Dorr-duplex  Classifier 

6  Hp. 


3 -Metering  Tanks 
2 -.Merrill  Presses 


-  5  x  6,  75  Hp. 

Allis-Chalmers  Ball  Mills 
2  -  Dorp-duplex 

Classifiers 


3-Hp. 
Erenier  Pump 


Triplex  Pumps 
1  -  10'x  12'Vac.  Pump 
Merrill  Zinc  Feeder 


8  -  Dorr  Thickeners 
Continuous  Counter  -  Current  Decantation 


27'jxfT/       \27'xJfX       \20jil£/        V  30  x  9 

FIG.  129.— Flow-sheet  of  Tom  Reed  Mill. 


tanks  the  solution  goes  to  the  barren  solution  tank  27  ft.  diameter  by  5  ft. 
deep.  The  barren  solution  is  piped  to  the  third  tank  of  each  decantation 
Series.^  Wash-water,  in  quantity  sufficient  to  replace  that  discharged  in 


THE  UNITED  EASTERN  MILL  209 

the  tailings,  is  added  in  the  fourth  tank  of  each  series.  For  refining  there 
are  provided  two  muffles  where  the  precipitate  is  dried  and  roasted  in  pans 
and  a  Steel  Harvey  tilting  furnace  Fig.  3,  where  the  roasted  precipitate 
is  melted  down  into  an  ingot. 

THE  UNITED  EASTERN  MILL,  OATMAN,  ARIZ.;  300  TONS  DAILY  CAPACITY 

The  ore  is  a  mixture  of  calcite  and  quartz  with  some  undecomposed 
andesite.  It  contains  1  oz.  gold  and  but  0.34  oz.  silver  per  ton.  The  gold 
is  so  finely  disseminated  that  fine  grinding  is  essential.  Figs.  130  and  131. 

Treatment. — This  consists  of  (1)  single-stage  coarse  crushing;  (2)  two- 
stage  ball-milling  in  cyanide  solution;  (3)  combined  air  and  mechanical 
agitation;  (4)  straight  counter-current  agitation;  (5)  removal  of  air  by  the 
Crowe  vacuum  treatment;  (6)  precipitation  by  zinc  dust  using  the  Merrill 
process;  (7)  fluxing  and  melting  the  wet  precipitate  into  bars. 

With  the  exception  of  the  Merrill  filter-press  for  separating  the  pre- 
cipitate no  filters  are  used  about  the  mill.  It 'is  remarkable  that  such  a 
high-grade  ore  can  be  well  extracted  by  counter-current  decantation. 

Coarse  or  Preliminary  Crushing. — All  lumps  of  ore  that  will  pass  a  10- 
in.  grizzly  are  sent  to  a  gyratory  crusher,  chosen  rather  for  a  sufficient- 
sized  jaw-opening  for  the  grizzly  discharge,  than  for  its  capacity  in  excess, 
this  being  35  tons  per  hour.  Lately  the  fines  of  this  run  of  mine  ore 
have  been  screened  out  by  grizzlies  set  at  If  in.  opening,  thus  relieving 
the  crusher  and  reducing  crusher  repairs.  It  is  now  thought  that  this 
crushing  might  better  be  performed  in  two  stages  for  the  production  of  a 
finer  product;  it  would  secure  a  finer  feed  for  the  ball-mills  that  come  next, 
thus  increasing  their  capacity. 

One  should  note  here  that  for  an  assured  supply,  storage  for  several 
days  should  be  provided.  This  point  is  often  overlooked. 

Coarse  Grinding. — From  the  feed  bin  the  ore  is  fed  to  the  ball-mill 
by  the  traveling  feeder,  Fig.  320.  The  setting  of  a  feed  gate  and  the 
varying  of  the  specific  gravity  of  the  pulp  in  a  special  Dorr  classifier 
determine  how  thick  a  feed  shall  come  out.  The  coarse  grinding  is  done  in 
a  Marcy  ball-mill  in  closed  circuit  with  a  Dorr  classifier  (see  Fig.  40), 
forged  chrome  iron  balls  being  used  in  the  mill.  The  grinding  is  done  in  a 
1.6  per  cent  KCN  solution,  showing  1  Ib.  protective  alkalinity. 

Fine  Grinding. — This  is  done  in  two  ball-mills  5  ft.  diameter  by  6  ft. 
long,  also  in  closed  circuit  with  Dorr  classifiers,  each  mill  being  of  90  tons' 
capacity  and  giving  a  product,  82  per  cent  of  which  is  minus  200-mesh. 
The  mills  both  in  the  coarse  and  fine  grinding  discharge  a  product  of  30 
per  cent  moisture  only. 

Agitation. — From  the  fine-grinding  department  the  pulp  flows  to  No.  1 
Dorr  thickener,  marked  U  on  the  flow-sheet,  Fig.  131,  arriving  there  with 


210 


TYPICAL  GOLD  MILL  PRACTICE 


a  specific  gravity  of  1.12  and  a  dilution  of  1  part  ore  to  4.5  parts  solution. 
The  thickened  underflow  from  this  thickener  of  a  specific  gravity  1.4  and 
amounting  to  290  tons  daily  is  pumped  to  the  seven  agitators  Nos.  1  to  7 


/A~~ 

i__j        ^-LBarren  Solution 


FIG.  130.— Plan  of  United  Eastern  Mill. 


of  the  plan  and  indicated  by  the  single  rectangle  A  of  the  flow-sheet, 
there  being  an  addition  of  197  Ib.  of  NaCN  to  the  first  agitator  as  this  is 
done.  The  period  of  agitation  is  sixty-two  hours  for  complete  solution. 


THE  UNITED  EASTERN  MILL 


211 


Agitation  is  carried  on  by  air  at  a  pressure  of  30  Ib.  per  square  inch,  using 
two  belt-driven  compressors  of  15  kw.  each. 

Thickening. — From  the  thickeners  A  the  pulp  flows  through  five  40-ft. 
thickeners  marked  No.  2  to  No.  6  on  the  flow-sheet  and  plan.  These  are 
arranged  for  straight  counter-current  work.  The  pregnant  solution  from 
No.  1  agitator  after  passing  the  zinc-boxes  becomes  barren  solution  to  go 
to  No.  4  thickener  and  the  wash  water  is  introduced  at  No.  6  thickener. 
The  flow  at  No.  6  is  split  and  about  one-third  is  sent  to  No.  7  thickener 
(see  the  plan).  The  underflow  discharges  of  Nos.  6  and  7  flow  together  to 
the  tailings  pond  at  a  moisture  content  of  0.82  ton  solution  per  ton  of 
ore. 

No.  1  or  primary  thickener,  which  takes  the  dilute  overflow  from  the 
last  Dorr  classifiers,  has  a  settling  area  of  4.37  sq.  ft.  per  ton  of  ore.  It  is 
necessary  to  operate  it  with  a  low  mud  line  to  prevent  colloidal  material 
getting  over  into  the  gold  solution  tanks.  About  870  tons  of  solution 
pass  to  the  press  solution  vat  while  600  tons  go  back  to  the  mill  storage. 


To  Tailings 


ItMi 


FIG.  131.— Flow-sheet  of  United  Eastern  Mill. 


The  regulation  of  the  thickeners  is  maintained  by  varying  the  speed  of 
the  diaphragm  pumps  which  feed  them,  the  specific  gravity  of  the  dis- 
charge, and  the  depth  of  the  mud  line  being  recorded  every  four  hours. 
In  maintaining  alkalinity  lime  is  added  dry  to  the  Marcy  mill  feed. 
The  loss  per  ton  is  $3.215  and  the  loss  of  cyanide  0.525  Ib.  per  ton  of  ore. 

Cost  of  Plant.— The  cost  of  the  mill  proper  is  $133,539.09,  to  which 
should  be  added  the  crushing  plant  $11,975.37.  the  coarse  ore  bins  $1,916.27, 
the  refinery  $5,789.13,  and  the  lime  house  $272.54,  or  a  total  of  $153,492.40, 
which  on  a  basis  of  200  tons  daily  would  be  $767.46  or  $2.13  per  annual  ton. 

Operating  costs  are  based  upon  $5  per  day  for  helpers,  $5.50  for  mill 
men  and  $6  for  solution  shift  foremen. 

In  1918  the  cost  of  operations  averaged  for  operating  labor  $0.47, 
repair  labor  $0.09,  supplies  $1.06,  power  $0.52,  miscellaneous  $0.04,  making 
a  total  of  $2.17  per  ton.  Coarse  crushing  cost  $0.07;  coarse  grinding  $0.32, 
and  fine  grinding  $0.50  per  ton. 

As  respects  particular  costs  also  in  1918  per  ton  of  ore:    Those  for 


212  TYPICAL  GOLD  MILL  PRACTICE 

heating  solution  were  $0.03,  of  general  expense  $0.18,  of  lighting  $0.008, 
of  assaying  $0.03,  sampling  $0.03,  of  tailing  disposal  $0.02,  of  clarification 
$0.04,  and  of  precipitation  $0.12. 

The  metallurgical  report  shows  in  1918  that  92,339  tons  were  milled 
carrying  100,903.02  of  gold  and  54,137.02  of  silver,  with  a  recovery  of 
96.75  Der  cent. 


CHAPTER  XVII 
TREATMENT  OF  GOLD  MILL  CONCENTRATES 

CLASSIFICATION 

For  this  treatment  it  is  recommended  that  for  the  grinding  a  tube-mill 
be  used  with  a  Dorr  classifier  in  closed  circuit.  It  yields  an  overflow,  which 
passes  on  for  agitation,  and  an  underflow  or  spigot  discharge,  which  goes 
back  to  the  tube-mill.  In  this  manner  no  granular  particles  can  escape 
the  grinding  action  of  the  tube-mill.  The  ground  product  goes  to 
Dorr  agitators  followed  by  Dorr  thickeners  to  be  subjected  to  continuous 
counter-current  agitation.  Even  then,  before  the  tailings  are  sent  to  waste, 
they  are  filtered  and  washed,  using  Oliver  or  similar  filters.  On  the  other 
hand,  the  clear  overflowing  pregnant  solution  of  full  strength  is  sent  to  the 
Merrill  precipitation  and  clarifying  process.  The  solution  is  handled  by 
aid  of  triplex  plunger-pumps.  For  the  above  methods  of  treatment  a 
side-hill  location  may  be  used,  with  ample  fall  throughout  with  good 
dumping  ground  below  for  the  tailings.  This  should  be  so  arranged  that 
the  tailings  can  be  stored,  not  carelessly  allowed  to  run  to  waste.  In  the 
planning,  the  all-gravity  arrangement  is  preferred  so  as  to  avoid  the  use  of 
troublesome  bucket  elevators  and  pumping.  The  raising  of  solutions  by 
pumping  is  another  matter,  even  to  raising  them  60  or  70  ft.  from  sump  to 
supply  tank;  there  is  no  grit  in  the  solution,  a  fruitful  cause  of  pump 
wear. 

The  concentrates  which  are  to  receive  the  special  treatment  above  spe- 
cified come  to  this  special  treatment  plant,  either  in  boxes  filled  from  the 
head  launder  of  the  tables  or  in  cars  that  have  directly  caught  the  product, 
or  indeed  into  a  launder  that  by  aid  of  a  stream  of  water  carries  it  to  the 
sump  of  a  sand  pump,  this  pump  lifting  it  to  collecting  vats  where  it  is 
drained  before  subjecting  to  the  special  treatment. 

The  stuff  can  be  treated  in  several  ways. 

(1)  It  may  be  treated  with  strong  cyanide  solutions  in  ordinary  perco- 
lating vats  for  an  extended  time. 

(2)  It  may  be  stacked  to  expose  it  to  the  weather,  then  percolating  it 
as  in  method  (1). 

(3)  It  may  be  shipped  at  once  to  the  smelting  works. 

(4)  It  may  be  treated  by  chlorination. 

213 


214        TREATMENT  OF  GOLD  MILL  CONCENTRATES 

(5)  It  may  be  all-slimed  as  outlined  above,  then  agitated  with  cyanide 
solution  and  the  tailings  carefully  filtered  and  washed. 

(6)  It  may  be  roasted,  then  finely  ground,  agitated  and  the  residue 
filtered. 

A  clean  iron  pyrite  may  be  treated  by  method  (1)  with  fair  results,  but 
time,  cyanide  consumption,  and  the  necessary  retreatment  are  against  it. 
At  a  small  mine  in  an  out-of-the-way  district,  as  a  temporary  method  it  is 
worth  trying.  -  j 

When  concentrate  to  be  treated  by  method  (2)  has  been  weathered, 
it  becomes  highly  acid,  and  needs  special  alkaline  washes. 

According  to  method  (3)  all  complex  concentrates,  that  is,  those 
containing  copper,  zinc,  or  lead,  or  complicating  impurities,  should  be 
smelted,  although  in  Western-Australia  pyrite  containing  as  much  as  21 
per  cent  arsenic  has  been  roasted  and  cyanided  with  a  90  per  cent  extraction. 

Chlorination  or  method  (4)  is  not  much  used  now.  At  Bendigo,  Vic- 
toria, Australia,  one  custom  "  pyrite  "  works  uses  chlorine  in  the  vats,  and 
this  pyrite  contains  as  much  as  12  per  cent  arsenic. 

Referring  to  method  (5)  the  concentrate  plant  of  the  Treadwell  group 
of  mines  uses  fine  grinding  and  agitation  with  cyanide  solution.  At  Waihi, 
Western  Australia,  a  concentrate  treatment  plant  specially  treated  5368 
tons  in  1911  equal  to  1.42  per  cent  of  the  ore  crushed  at  the  mills.  This 
concentrate  product  contained  5.25  oz.  in  gold  and  3.2  oz.  of  silver  per 
ton,  and  the  recoveries  were  96.2  per  cent  of  the  gold  and  94.8  per  cent 
of  the  silver,  at  a  cost  of  $6  per  ton  of  concentrates  treated. 

As  to  method  (6),  at  many  places  the  concentrate  is  roasted  prior  to 
cyanidation,  especially  at  the  Goldfield  Consolidated,  Goldfield,  Nevada, 
and  at  Kalgoorlie,  Western  Australia.  At  the  Ivanhoe  mine  at  the 
latter  place,  the  annual  output  to  be  treated  is  24,000  tons,  which  is  col- 
lected in  ordinary  tanks  roasted  in  five  Edwards  furnaces,  Fig.  74,  mixed 
with  cyanide  solution,  ground  in  five  5-ft.  pans.  Fig.  135,  agitated  and  filter- 
pressed.  An  extraction  of  95  per  cent  was  obtained  at  a  cost  of  $2.68 
per  ton. 

The  five  mills  on  Douglas  Island,  Alaska,  contain  a  total  of  900  stamps, 
and  crush  approximately  5000  tons  per  day.  The  crushed  ore  after  amal- 
gamation is  concentrated  on  360  Frue  vanners,  yielding  an  average*  of  90 
tons  of  concentrate  per  day,  of  from  2.5  to  4  oz.  gold  per  ton.  To  treat 
this  product,  a  cyanide  mill,  owned  jointly  by  the  Alaska-Tread  well  Gold 
Mining  Company,  the  Alaska  Mexican  Gold  Mining  Company,  and  the 
Alaska  United  Gold  Mining  Company,  was  built  to  treat  the  concentrate 
made  in  their  various  mills,  and  has  proved  an  unqualified  success. 


THE  ALASKA-TREADWELL  MILL 


215 


THE  ALASKA-TREADWELL  CONCENTRATE  TREATMENT-PLANT 

consists  of  three  buildings  situated  on  a  hillside  200  ft.  above  the  stamp- 
mill.  The  upper  building  contains  the  grinding  and  amalgamating  plant, 
with  a  lower  floor  for  solution-storage  tanks.  The  lower  contains  the 
cyanide  equipment  proper,  while  the  refinery  is  in  a  concrete  building  at 
one  side.  The  flow-sheet  of  operations  is  shown  in  Fig.  133. 

The  concentrate  is  received  in  two  100-ton  steel  storage  bins,  4,  4, 
15  ft.  diameter,  with  55°  conical  bottoms  Here  it  is  kept  covered  with 
water,  which  effectually  prevents  oxidation  of  the  sulphides.  From  this 
point  until  the  cyanide  treatment  begins,  the  concentrate  is  in  strong  lime 
solution  at  all  tunes.  At  the  apex  of  the  conical  bottom  of  each  bin,  tight- 
fitting  gates  control  the  outflow,  which  is  at  once  sluiced  directly  into  Dorr 


FIG.  132. — Outline  of  Tube  Mill  Circuit,  Alaska-Treadwell  Mill. 

classifiers,  5,  5.  The  sluicing  medium  is  the  coarse  return  product  referred 
to  later.  There  are  two  Dorr  classifiers  driven  by  one  7.5  H.P.  electric 
motor,  one  feeding  into  each  tube-mill  and  making  twenty-four  strokes  per 
minute.  This  rate  of  speed  (causing  greater  agitation)  was  found  neces- 
sary to  separate  the  large  bulk  of  the  fine  from  the  coarse. 

The  coarse  product  of  the  classifiers  falls  into  the  spiral  feeders  of  the 
tube-mills.  These  mills,  6,  6,  are  of  the  Abbe  type,  5  by  22  ft.,  with  cor- 
rugated sectional  liners:  and  3-in.  Danish  flint  pebbles,  are  used  for  the 
grinding.  Amalgamation  was  formerly  part  of  the  process,  but  the  whole 
product  is  now  being  cyanided  direct  without  this.  From  a  sump  in  the 
launder,  an  air-lift  elevates  the  pulp  to  a  spitzlutte,  from  which  the  coarse 
material  is  continuously  drawn  into  a  Dorr  classifier,  11,  the  coarse  from 
which  feeds  a  4  by  12-ft.  Abbe  tube-mill,  12,  similar  to  the  larger  ones 


216 


TREATMENT  OF  GOLD  MILL  CONCENTRATES 


described  above.  The  discharge  from  this  mill  joins  the  overflow  from 
the  spitzlutte,  and  is  elevated  by  air-lifts  to  two  settling-cones,  so  situated 
that  the  spigot-discharge  from  them  becomes  the  sluicing  medium  for  the 
original  feed  referred  to  above.  The  overflow  from  the  Dorr  classifiers 

passes  into  two  Callow  dewatering  cones, 
the  spigot  product  of  which  flows  into 
launders,  thence  into  a  6-in.  pipe,  37  ft. 
long,  having  a  fall  of  f  in.  per  foot,  which 
conveys  the  pulp  directly  to  the  lower  or 
cyanide  building.  In  the  lower  building 
the  pulp  is  received  into  a  wooden  dis- 
tributing-box, from  which  it  flows  into  four 
8-ft.  Callow  cones.  The  spigot-product 
from  these  cones  discharges  into  four 
similar  ones  placed  lower  than  the  first 
set. 

The    spigot-product    from   the   lower 
cones  enters  one  of   four   Pachuca  tanks, 
22,  where  it  receives  a  preliminary  treat- 
ment of  three  hours'  agitation  in  a  solution 
containing  2  Ib.  of  lime  per  ton  (0.1  per 
cent),  after  which  it  is  allowed  to  settle  and 
the  clear  solution  is  decanted.     The  filling, 
agitating,  settling,  decanting,  and  discharg- 
ing of    a    25-ton  charge   of    concentrate, 
which  includes  46  tons  of  lime  solution, 
requires  somewhat  less  than  twenty-four 
hours.     This  preliminary  treatment  saves 
in  the  subsequent  treatment  at  least  1  Ib. 
of  cyanide  per  ton  of  concentrate.     The 
overflow  lime  water  from  the  Callow  cones 
enters  the  same  sump  with  the  decanted 
lime-water    from    the    prelimi- 
nary treatment,  and  is  pumped 
into    a  reservoir    of    1$    tons' 
™ks-    capacity  situated  in  the  upper 
Alaska  Treadweil  Mill,  building.     The  thickened  pulp, 
ranging  from  1.8  to  2.2  specific 

gravity,  is  drawn  into  one  of  eight  Pachuca  agitation  tanks,  24,  where 
it  is  given  the  cyanide  treatment.  All  Pachuca  tanks  in  the  mill  are  10  ft.  in 
diameter  and  30  ft.  high,  with  60°  conical  bottoms.  When  filled  to  the 
level  found  best  for  agitating  (which  is  6  in.  below  the  top  of  the  central 
column),  each  tank  holds  a  volume  equivalent  to  50  tons  of  water. 


Bullion  to  mint. 
Slag  to  smelter. 
Amalgam. 
Solution  returned 


FIG.  133.— Flow-sheet. 


THE  ALASKA-TREADWELL  MILL  217 

This  is  equal  to  the  regular  charge  of  30  tons  of  concentrate  with  40  tons 
of  solution.  The  floors  under  the  Pachuca  tanks,  as  well  as  all  other 
floors  in  the  building,  are  of  smooth  concrete,  sloping  to  a  central  sump, 
suppled  with  small  pumps  to  return  any  escaped  solution  and  to  pump 
it  to  the  proper  tanks. 

The  first  cyanide  treatment  consists  of  eight  hours'  agitation  in  a  2-lb. 
(0.1  per  cent)  cyanide  solution,  either  potassium  or  the  mixed  cyanides 
being  successfully  used.  Alkali  is  kept  at  1.25  Ib.  (0.063  per  cent)  of  lime 
(CaO)  per  ton  of  solution.  Lime  is  added  during  the  treatment  if  the 
titrations  show  below  that  figure;  eighteen  hours  is  allowed  for  settlement 
and  decantation  of  this  solution.  Decantation  takes  place  through  a 
flexible  hose. 

The  long  settlement  allowed,  with  the  excessively  fine  condition  of  the 
concentrate,  its  high  specific  gravity,  from  4.6  to  5.0,  and  the  high  alka- 
linity of  the  solution,  leaves  a  30-ton  packed  mass  in  the  bottom  of  the 
Pachuca.  This  is  brought  into  agitation  within  fifteen  minutes  by  a 
device  designated  as  the  "  spider,"  which  is  an  adjustable  hollow  annular 
casting  with  radiating  fingers,  the  whole  encircling  the  central  agitation- 
column,  see  Fig.  100. 

The  second  cyanide  treatment  of  the  charges  is  with  solution  drawn 
from  the  barren-solution  storage  tanks  or  the  wash-solution  storage,  the 
cyanide  strength  being  1.5  Ib.  (0.075  per  cent)  per  ton  of  solution.  After 
two  hours'  agitation  the  air  is  shut  off  and  almost  immediately  decantation 
is  started.  This  decanted  solution  is  pumped  directly  on  to  an  incoming 
fresh  charge,  being  strengthened  in  cyanide  as  it  enters  the  tank,  and 
becoming  the  first  cyanide  solution  for  the  new  charge.  This  cycle  in 
handling  solution — barren  to  wash-solution,  then  to  second  cyanide  treat- 
ment at  0.075  per  cent  cyanide,  then  to  first  treatment  at  0.1  per  cent 
cyanide,  thence  to  precipitation  and  back  to  barren — gives  at  each  step, 
just  the  conditions  best  suited  for  that  step,  and  is  very  satisfactory  in 
practical  operation.  The  settled  pulp  after  the  second  decantation  has  a 
specific  gravity  of  1.8,  and  is  readily  agitated  by  means  of  the  spider,  and 
then  discharged  into  the  pulp-storage  tank  by  a  Byron  Jackson  4-in. 
centrifugal  pump. 

Filtering  is  done  in  two  type-l-B  Kelly  presses.  By  opening  valves 
in  the  circulation-lines  directly  under  each  press  it  is  filled  with  either 
pulp  or  wash-solution  as  desired.  The  excess  pulp  or  wash-solution  from 
the  press-cylinder  is  returned  into  its  proper  line  by  displacing  with  com- 
pressed air  admitted  into  the  cylinder.  The  amount  of  wash  given  depends 
upon  the  comminution  of  the  concentrate,  the  usual  pulp  being  washed 
with  0.5  ton  of  solution  per  ton  of  concentrate.  The  cake  formed  during 
decantation  of  the  first  treatment-solution,  being  very  fine  slime  and  more 
impervious  to  wash-solution  than  the  regular  pulp,  is  given  1  ton  of  wash 


218  TREATMENT  OF  GOLD  MILL  CONCENTRATES 

per  ton  of  concentrate.  When  filling  the  press,  the  contained  air  is  allowed 
to  escape  through  an  overhead  pipe  attached  to  the  highest  point  of  the 
press-cylinder.  The  change  in  sound  of  the  exhaust  indicates  to  the 
pressman  when  the  press  is  full.  After  drying  the  cake  with  compressed 
air  until  it  contains  not  more  than  10  per  cent  of  moisture,  the  press  is 
opened  and  the  cakes  shaken  off  with  wooden  paddles,  and  then  sluiced 
with  water  to  the  tailing-dam.  A  distributer  below  the  press-launder  sends 
the  gold-solution  to  two  gold-sumps  and  the  wash-solution  to  the  two 
wash-solution  storage-tanks.  These  four  tanks,  as  well  as  a  clarifying- 
tank  which  is  in  the  same  group,  are  built  of  3-in.  redwood,  15  ft.  in  diam- 
eter by  16  ft.  deep,  and  each  holds  75  tons  of  solution. 

The  wash-solution  is  pumped  to  a  Pachuca  tank  as  needed,  becoming  a 
second-treatment  solution.  From  the  gold-tank  the  solution  is  drawn 
into  the  clarifying-tank,  in  which  are  suspended  vertically  six  canvas  filter- 
leaves,  all  connected  to  the  suction  of  a  triplex  7  by  9-in.  Aldrich  electric 
pump,  used  exclusively  for  pumping  gold-solution  through  the  precipita- 
tion-presses. A  traveling-belt,  driven  by  ratchet-gears  and  a  pair  of 
eccentrics  connected  to  the  pump-drive,  feeds  zinc-dust  into  a  cone.  Here 
the  dust  is  emulsified  with  a  small  stream  of  gold-solution  tapped  from  the 
discharge-column  of  the  same  pump,  and  is  then  drawn  into  the  suction- 
line.  An  automatic  float  in  the  cone  prevents  the  introduction  of  air  into 
the  pump-suction.  The  pump  raises  the  solution  with  the  zinc  dust  to 
the  upper  part  of  the  building  and  forces  it  through  two  36-in.  triangular, 
16-frame  Merrill  presses.  An  average  of  145  tons  of  solution  is  precipi- 
tated daily,  with  a  consumption  of  1-3  Ib.  of  zinc  dust  per  ton  of  solution, 
equivalent  to  0.86  Ib.  of  zinc  dust  per  ton  of  concentrate.  The  average 
strength  of  solution  before  precipitation  is  1.25  Ib.  (0.0625  per  cent)  of 
cyanide;  1  Ib.  (0.05  per  cent)  of  lime,  and  $9.50  (9.2  dwt.)  gold.  The 
barren  or  precipitated  solutions  are  kept'  at  10  cents  (2.3  grains),  or  less, 
gold  per  ton,  and  are  used  for  wash-solution  or  returned  to  the  Pachuca 
tanks,  as  desired. 

CONCENTRATE-TREATMENT   AT  THE    GOLDFIELD    CONSOLIDATED    MILL 

GOLDFIELD,  NEV. 

The  raw  concentrate  amounting  to  6  per  cent  of  the  weight  knd  con- 
taining 67  per  cent  of  the  value  of  the  ore,  is  collected  in  flat-bottomed- 
agitator  tanks.  It  is  here  neutralized  with  lime  and  pumped  to  three 
Pachuca  agitators,  in  which  it  is  agitated  during  eight-hour  periods  in  a 
2-lb.  solution  of  cyanide.  Decantation  at  the  end  of  the  period  is  still  prac- 
ticed and  the  charge  is  re-agitated  with  a  freshly  precipitated  solution. 
Five  periods  of  eight  hours  each,  followed  by  decantation,  are  sufficient 
to  remove  from  80  to  85  per  cent  of  the  value  of  the  concentrate.  It  is  the 
intention  to  send  to  the  roaster  a  product  valued  at  $25  to  $30,  and  this 


GOLDFIELD  CONG.  MILL  219 

treatment  is  varied  with  the  grade  of  the  ore  so  as  to  accomplish  this  result. 
The  pulp  from  the  Pachucas,  when  dissolution  is  completed,  is  delivered 
to  a  storage  tank  from  which  it  is  pumped  to  Kelly  filter-press  for  fil- 
tration and  drying.  This  drying  is  accomplished  with  air  and  the  moisture 
is  reduced  to  12  per  cent.  The  consumption  of  cyanide  during  the  raw 
treatment  is  2.5  Ib.  per  ton  and  lead  acetate  is  used  in  the  proportion  of 
1  Ib.  per  ton  of  concentrate. 

The  product  is  dumped  into  a  bin,  and  a  14-in.  conveyor,  set  at  an  angle 
of  17°,  carries  it  over  a  Blake-Dennison  automatic  weighing  machine  en 
route  to  the  bins  in  the  roasting  plant.  The  concentrate  from  this  con- 
veyor is  distributed  by  means  of  a  swinging  bucket  elevator  to  two  bins 
having  45°  sloping  bottoms  and  1620-cu.  ft.  capacity,  from  which  it  is  fed 
by  means  of  two  12-in.  screw-conveyors,  making  three-quarters  of  a  revo- 
lution per  minute,  to  two  slow-moving  belts.  These  belts  discharge  the 
concentrate  through  the  arches  of  the  furnace  between  the  first  two  rabbles. 

Roasting  of  concentrates  is  done  in  two  54-spindle  duplex  Edwards 
furnaces,  each  having  1456  sq.  ft.  hearth  area.  The  capacity  of  each 
furnace  is  40  tons  of  concentrates  per  day,  although  the  amount  roasted 
in  the  two  furnaces  is  approximately  55  tons  per  day.  The  raw  concen- 
trates, after  a  preliminary  cyanide  treatment,  assay  1.23  oz.  Au  and  18.76 
per  cent  S,  the  sulphur,  after  roasting,  being  reduced  to  0.90  per  cent. 
The  cost  of  roasting,  per  ton  of  concentrate,  is  $0.82;  while  the  complete 
cost  of  the  two  furnaces,  dust  flues,  stack,  65  by  158  ft.  steel  building 
and  miscellaneous  bins  and  machinery  amounted  to  $70,459.16.  The  con- 
centrate loses  17  per  cent  of  its  weight  in  roasting;  and  of  the  1J  per  cent 
of  the  material  passing  out  of  the  furnace  as  dust,  only  J  per  cent  is  lost 
out  of  the  stack.  Five  Merton  furnaces  for  the  roasting  of  120  tons  per 
day  of  Kalgoorlie  sulpho-telluride  ore  cost  $38,900. 

The  bins  are  filled  on  the  day  shift,  and  have  sufficient  capacity  to  run 
for  twenty-four  hours.  Each  furnace  requires  4J  H.P.  By  means  of  iron 
goose-neck  flues,  the  gases  from  the  roasters  at  a  temperature  of  450°  F. 
are  delivered  to  a  concrete  dust-flue  264  ft.  long,  having  a  cross-section  of 
50  sq.  ft.  From  this  flue,  20,700  cu.  ft.  of  gases  per  minute  escape  through  a 
steel  stack,  100  ft.  high  and  54  in.  diameter  having  a  temperature  at  the 
base  of  the  stack  of  325°  F.  Velocity  of  the  gases  in  the  dust-flue  is 
1\  ft.  per  second. 

The  roasted  ore  is  discharged  into  a  Baker  cooler,  5  by  22  ft., 
revolving  in  water  with  about  40  per  cent  submergence.  It  is  delivered 
to  one  tank  for  twenty-four  hours,  then  settled  and  decanted  to  a  consistence 
of  1  to  1  and  sulphuric  acid  added  in  the  proportion  of  20  Ib.  per  ton 
of  concentrate.  Agitation  with  the  sulphuric  acid  is  continued  for  eight 
hours.  Water  is  then  added  to  fill  the  tank  and  the  charge  allowed  to 
settle.  When  clear,  the  wash  is  decanted  and  the  tank  refilled  with  fresh 


220        TREATMENT  OF  GOLD  MILL  CONCENTRATES 

water.  Four  water  washes  are  given,  equivalent  to  eight  tons  of  wash 
water  per  ton  of  concentrate.  All  washes  are  clarified  and  the  overflow 
sent  to  six  redwood  tanks,  10  ft.  diameter  and  5  ft.  high,  arranged  in 
series  for  recovering  the  copper.  These  tanks  are  kept  filled  with  cyanide 
tins  and  all  kinds  of  scrap  from  the  mill.  The  average  copper  content  of 
the  washes  is  0.4  Ib.  per  ton,  and  70  per  cent  is  recovered. 

The  thoroughly  washed  charge  is  neutralized  with  lime,  and  by  means 
of  centrifugal  pumps  elevated  to  one  of  four  Pacfruca  agitators,  14  ft. 
diameter  by  25J  ft.  high.  Here  the  roasted  charge  is  agitated  for  eight 
hours  in  a  2-lb.  solution  of  cyanide,  containing  1.2  Ib.  CaO  as  protective 
alkali.  At  the  end  of  eight  hours,  agitation  is  discontinued,  the  charge 
settled,  decanted,  and  re-agitated  with  a  freshly  precipitated  solution 
in  the  same  manner  as  described  above  in  the  treatment  of  the  raw  con- 
centrate. Five  periods  of  agitation  followed  by  decantation  are  given,  and 
a  total  of  3  tons  of  solution  per  ton  of  concentrate  is  decanted.  Consump- 
tion of  chemicals  amounts  to  4J  Ib.  cyanide  and  2  Ib.  lead  acetate  per  ton 
of  concentrate.  After  agitation  is  completed,  the  settled  charge  is  deliv- 
ered to  a  storage  tank  18  ft.  diameter  by  8  ft.  high,  fitted  with  the  adjust- 
able square-shaft  agitator.  Placed  centrally  in  the  bottom  of  this  tank 
is  a  4-ft.  cone  with  pipe  connections  through  which  the  thickened  pulp  is  fed 
to  a  5  by  18-ft.  tube-mill.  The  pulp  issuing  from  the  tube-mill  is  elevated 
by  means  of  a  belt  and  bucket  elevator  back  to  the  above-mentioned  storage 
tank.  This  circulation  grinding  is  continued  for  sixteen  hours,  at  the  end 
of  which  time  95  per  cent  of  the  material  will  pass  a  200-mesh  screen. 
From  80  cents  to  $1.25  per  ton  is  removed  in  this  circuit.  Since  the  change 
of  solution  increases  extraction  and  since  the  final  tailing  is  sent  to  the  mill 
proper  for  filtration,  it  was  decided  to  re-grind  after  the  greater  part  of  the 
gold  had  been  removed.  After  re-grinding,  the  pump  is  delivered  by 
means  of  a  centrifugal  pulp  to  the  filter  storage  tank  in  the  mill  proper, 
mixed  with  the  mill  pulp,  filtered,  and  sent  to  waste.  Costs  are  as  follows  : 
Labor,  $1.02;  power,  $0.78;  and  total  supplies,  $3.88;  making  a  total 
cost  of  $5.68  per  ton.  During  1912  results  were  as  follows:  value  of  raw 
concentrate,  6.58  oz.  gold;  value  after  treatment,  1.23  oz.  gold;  recovery, 
81.3  per  cent;  value  of  tailing  after  roasting  and  treating  0.097  oz.  gold; 
recovery,  92.16  per  cent;  total  recovered  from  roasted  material,  1?.23  per 
cent;  and  recovered  from  both  treatments,  98.53  per  cent. 

The  treatment  at  each  plant  was  as  follows: 

(A)  Raw  concentrates  are  agitated  in  2  Ib.  KCN  solution  for  five  eight- 
hour  periods,  decanting  after  each  period;  after  filter-pressing,  they  'are 
roasted  in  two  Edwards  duplex  54-spindle  furnaces.  The  roasted  concen- 
trates are  agitated  for  eight  hours-  in  H2SO4  solution  (20  Ib.  acid  per  ton  of 
concentrates).  After  four  water  washes,  the  charge  is  neutralized  ard  agi- 
tated in  2  Ibs.  KCN  for  five  eight-hour  periods,  decanting  after  each. 


CYANIDING  CONCENTRATES 
CYANIDATION  OF  GOLD  BEARING  CONCENTRATES 


221 


ASSAY  or. 

PER   TON. 

EXTRACTION 

COST  PER  TON  OF  CONC'TS. 

a 

8 

.2 

Plant. 

I 

if 

3 

C 

C 

1 

&£ 

fc 

1° 

f 

1 

o 
•3 

1 

1 

|| 

1 

1 

1 

n* 

• 

11 

1 

Au. 

Ag. 

Au. 

Ag. 

Ibs. 

0 

H 

Q 

^ 

i 

£ 

1 

A—  Goldfield, 

Cons.,  Nev  

55 

6  58 

98.53 

7. 

.33 

.82 

2.34 

.77 

.30 

.18 

$5.6£ 

B—  Alaska,  Tread- 

well  

75 

292 

96.5 

2.3 

.57 

.78 

.22 

.81 

2.81 

C  —  Geldenhuis 

Deep,  Transvaal 

16.0 

98.75 

3.3 

3.65 

2.67 

6.3'J 

D—  Oriental, 

Cons,  Korea.  .  . 

2.2 

93 

2.87 

Then  it  is  tube-milled  for  sixteen  hours  and  sent  to  the  regular  plant  filter, 
using  3  Ib.  lead  acetate  per  ton  of  concentrates.  (J.  W.  Hutchinson, 
Min.  and  Scientific  Press,  January  25;  February  1,  1913.) 

(B)  Clean  and  docile  pyritic  concentrate  is  ground  through  200  mesh; 
agitated  in  1.5  Ib.  KCN  for  eight  hours;  decanted;  agitated  for  four  hours 
with  barren  solution;   filter-pressed  and  precipitated  by  zinc  dust.     (1911 
Mine  Report,  Min.  and  Scientific  Press,  June  29,  1912.) 

(C)  Mill  scrap  and  black  sands.     After  amalgamation,  the  concentrate 
pulp  for  cyanidation  (98  per  cent  200  mesh)  assays  16  oz.  Au.     (R.  Lind- 
say, J.  C.  M.  &  M.  So.,  S.  A.) 

(D)  All-sliming  in  a  tube-mill  with  KCN,  then  forty-eight  hours'  com- 
bined air  agitation  and  leaching,  followed  by  filter-pressing. 

Aeration.— Oxygen-absorbing  compounds,  such  as  pyrrhotite,  horn- 
blende, etc.,  not  only  cause  an  increased  cyanide  consumption,  but  fre- 
quently reduce  the  extraction  of  the  gold  by  robbing  the  solution  of  its 
oxygen.  Suitable  oxidation  of  the  pulp,  prior  to  the  application  of  cyanide 
solution,  by  forcing  compressed  air  through  the  charge,  renders  many  of 
these  oxide-consuming  compounds  harmless. 

Roasting. — In  the  treatment  of  ore  in  which  the  gold  is  intimately 
associated,  either  physically  or  chemically,  with  such  telluride  compounds 
as  sylvanite  or  calaverite,  or  with  arsenical  or  antimony  compounds,  pre- 
liminary roasting  is  frequently  the  only  known  recourse.  The  practical 
object  of  the  roast  is  to  so  liberate  the  gold  as  to  permit  ample  contact 
with  the  cyanide  solution  and  destroy  deoxidizers  and  cyanicides.  (A 
"  dead  "  or  "  sweet  "  roast  is  usually  essential.)  The  most  extensive  appli- 
cation of  roasting  is  at  Kalgoorlie,  though  some  Cripple  Creek  ores  receive 


222  TREATMENT  OF  GOLD  MILL  CONCENTRATES 

this  preliminary  treatment,  as  do  the  graphitic  ores  of  the  Ashanti  Gold- 
fields.  In  the  treatment  of  rebellious  concentrates  by  cyanide,  roasting 
plays  an  important  part  at  the  Goldfield,  Nevada,  plant. 

The  usual  furnaces  are  the  Edwards,  Merton,  Pearce,  and  Holthoff. 

On  the  telluride  ores  of  Kalgoorlie,  crushed  through  approximately  28- 
mesh  roasting  with  Edwards  and  Merton  furnaces,  costs  approximate 
65  cents  per  ton  of  ore.  The  sulphur  content  of  the  raw  ores  varies  from 
3  to  6  per  cent,  although  the  elimination  of  the  sulphur  affords  only  a 
rough  indication  of  the  success  of  the  roast.  The  hearth  area,  per  ton 
of  ore  roasted  per  day,  varies  from  17  to  29  sq.  ft.  Consumption  of 
wood  varies  from  10  to  13  per  cent  of  the  weight  of  the  ore.  The 
temperature  is  approximately  650°  C. 


CHAPTER  XVIII 
VARIOUS   TREATMENTS    AND    CALCULATIONS 

FLOTATION  AND  CYANIDING  AND   CALCULATIONS 

Concentration  by  Flotation,  Cyaniding  or  Smelting  the  Concentrates 
and  Cyaniding  the  Tailings. — The  method  would  be  suited  to  a  low- 
grade  gold  ore,  that,  owing  to  its  refractory  nature,  could  not  be  profitably 
treated  by  cyaniding. 

First  Case. — The  ore  would  be  all-slimed,  then  subjected  to  flotation,; 
yielding  a  small  proportion  of  high-grade  flotation  concentrate  and  a  clean 
tailing  for  cyaniding.  The  concentrate  would  be  subjected  to  special 
cyanide  treatment  as  described  under  head  of  "  treatment  of  gold  mill 
concentrate." 

Second  Case. — The  ore  would  contain  copper  or  zinc  sulphides  that 
would  interfere  with  cyaniding.  The  ore  would  be  crushed,  saving  the 
concentrate  and  wasting  the  tailings.  The  concentrates  would  then  be 
shipped  to  the  smelter. 

Third  Case. — The  ore  could  be  cyanided  and  the  tailings  concentrated. 
In  this  case  the  aim  is  to  treat  by  cyaniding  alone,  but  from  the  tailings  to 
recover  some  concentrates  by  means  of  flotation.  The  concentrate  would 
be  subjected  to  special  cyanide  treatment. 

DRYING  AND  CYANIDING 

The  ore  of  moderate  grade  contains  graphite  that  would  interfere  with 
cyaniding.  The  crushed  ore  would  be  dried  at  150°  C.  so  as  to  render  the 
graphite  inactive  and  less  flocculent.  It  would  then  be  cyanided. 

TREATMENT  OF  TAILINGS  FROM  ACID  OR  AMMONIA  LEACHING 

The  original  ore  has  a  quartz  gangue.  If  treated  by  sulphuric  acid 
leaching  it  must  have  but  little  lime  or  magnesia  carbonates.  If  treated 
by  ammonia  those  carbonates  do  not  interfere.  In  either  case  the  ore 
carries  oxidized  copper  minerals  and  even  some  microscopic  metallic 
copper.  The  copper  having  been  removed  the  tailing  is  in  fair  state  to 
recover  any  contained  gold  or  silver  by  cyaniding  (see  "  Sulphuric  Acid  or 
Ammonia  Leaching  of  Copper  Ores.") 

CALCULATION  OF  TONNAGES  IN  MILLS 

In  wet-crushing  mills,  concentration  and  hydro-metallurgical  works  it 
is  often  desirable  to  measure  the  water  and  ore  handled,  either  the  amounts 

223 


224  VARIOUS  TREATMENTS  AND  CALCULATIONS 

contained  in  tanks  or  the  quantities  passing  in  a  given  time.  In  some  mills 
such  measurements  are  systematically  made,  but  in  others  the  amounts 
are  merely  guessed,  or  they  are  measured  once  and  ever  after  assumed  to 
remain  constant.  Discrepancies  between  theoretical  and  actual  recovery 
are  due  to  errors  in  sampling  and  assaying  the  material  before  treatment, 
added  to  the  corresponding  errors  affecting  the  material  after  treatment, 
and  multiplied  by  errors  in  the  estimate  of  the  tonnage  treated.  The 
last  item  is  therefore  fully  as  important  as  the  others  in  calculating  probable 
returns. 

The  tonnage  of  sand  in  vats  filled  by  settling  under  water  is  best  ascer- 
tained by  means  of  boxes  of  stout  sheet  iron  (conveniently  made  of  exactly 
1  cu.  ft.  capacity,  but  in  any  case  accurately  measured),  having  a  number  of 
small  perforations  in  the  bottom  and  provided  with  handles.  Several  of 
these  are  placed  in  the  vat  at  various  stages  of  the  filling,  and  are  allowed 
to  remain  throughout  the  treatment.  While  the  vat  is  being  discharged 
these  are  carefully  removed  and  "  struck  "  level;  the  contents  are  then 
dried  and  weighed,  giving  the  pounds  of  dry  solid  per  cubic  foot.  Several 
charges  should  be  thus  tested  and  averaged  to  obtain  a  constant  value  for 
the  ore  or  tailing  treated.  It  is  desirable  to  place  some  of  these  boxes  near 
the  center  and  others  near  the  periphery  of  the  vat,  so  as  to  represent  vari- 
ations in  horizontal  as  well  as  vertical  distribution.  The  mean  weight  per 
cubic  foot  and  the  volume  of  sand  in  the  charge  give  the  total  weight  of 
sand.  While  cubic  boxes  are  often  used,  a  cylindrical  form  is  preferable, 
as  being  less  liable  to  deformation. 

In  estimating  the  cubic  content  of  a  round  vat  several  diameters 
should  be  measured  (preferably  three  or  four  making  equal  angles  with 
each  other  at  about  the  middle  depth),  and  for  the  greatest  accuracy  a 
similar  set  of  measurements  should  be  made  near  the  bottom  and  another 
near  the  top,  the  arithmetic  mean  of  all  the  diameters  measured  being 
used  in  the  computation.  In  supposedly  cylindrical  wooden  vats  of  over 
25  ft.  diameter  differences  of  6  in.  or  more  may  be  found,  due  to  imperfect 
construction,  to  settling,  or  to  unequal  shrinkage  of  the  staves,  owing  to 
their  upper  portions  being  intermittently  dried  while  the  lower  ends  remain 
wet. 

If  D  be  the  internal  diameter,  and  H  the  depth,  in  feet,  of  a  cylindrical 
tank,  the  volume  is  0.7854  D2H  cu.  ft.,  0.024544  D2H  fluid  ton*s,  5.89 
D2H  U.  S.  gallons,  or  4.008  D2H  imperial  gallons.1 

1  The  U.  S.  gallon  will  hold  8£  Ib.  of  water,  the  British  imperial  gallon  10  Ib.  of  water. 
The  percentage  P,  of  dry  slime  in  the  pulp,  is  computed  by  the  formula 


P 


where  S  is  the  sp.  gr.  of  the  dry  slime,  and  a  the  sp.  gr.  of  the  wet  pulp.     The  sp.  gr.  of 
the  pulp  can  be  ascertained  by  a  hydrometer  or  by  weighing  a  unit  measure  of  it. 


TONNAGE  CALCULATIONS  225 

The  capacity  of  a  filter-press  is  most  accurately  determined  by  blowing 
the  charge  as  nearly  dry  as  possible  before  opening,  then  selecting  a  certain 
proportion  of  the  frames  at  equal  distances  from  end  to  end  of  the  press, 
weighing  the  entire  content  of  each  separately,  and  taking  an  individual 
moisture  sample  from  each  frame  tested.  The  dry  weight  of  slime  in 
each  is  separately  calculated  and  the  average  multiplied  by  the  number  of 
frames. 

In  ascertaining  the  weight  of  solid  in  a  vat  filled  with  uniformly  liquid 
pulp,  such  as  slime  in  an  agitator,  a  cubic  foot  or  any  convenient  measured 
volume  may  be  dried  and  the  residue  weighed,  whence  the  weight  in  the 
entire  volume  is  obtained  by  proportion.  If  the  mixture  is  weighed  before 
drying  the  percentage  of  solid  in  the  pulp  may  also  be  found. 

A  much  easier  and  more  rapid  method  is  to  find  the  specific  gravity  of 
the  pulp,  either  with  a  hydrometer  or  by  weighing  a  liter  or  other  conveni- 
ent volume.  The  density  of  the  dry  solid  must  also  be  known,  at  least 
approximately.  Knowing  these  two  values,  specific  gravity  of  mixture 
and  density  of  dry  solid,  the  weight  of  dry  solid  per  cubic  foot  can  at  once 
be  calculated. 


CHAPTER  XIX 
SMELTING  OF  GOLD  ORES 

BLAST-FURNACE  SMELTING  VS.  CYANIDING  OF  GOLD  ORES 

Gold  may  be  recovered  from  its  ore  by  the  processes  of  silver-lead,  or  of 
copper-matte  smelting.  It  often  is  found  in  copper-  or  lead-bearing  ores 
and  when  in  excess  of  0.02  oz.  per  ton,  is  paid  for  by  the  smelting  works  at 
the  rate  of  $19  to  $19.50  per  ounce.  Practically  all  the  gold  is  recovered 
in  smelting,  and  this  would  be  the  best  method  of  treatment  were  it  not 
for  the  high  cost  of  freight  and  for  treatment.  If  smelted  near  the  mine 
in  a  works  operated  by  the  mining  company,  the  cost  of  freight  is  elim- 
inated. The  charge  for  lead-free,  fairly  silicious  ores,  from  Cripple  Creek 
and  from  Boulder  county,  Colorado,  is  from  $4  to  $10  per  ton,  according 
to  grade.  The  low-grade  ores  are  subject  to  a  low-treatment  rate.  On 
the  other  hand,  ore  treated  by  milling  and  amalgamation,  or  by  cyanida- 
tion,  while  the  extraction  is  less,  often  yields  higher  net  returns.  A  sample 
is  found  in  the  case  of  the  silicious  gold  ore  from  Boulder  county,  Colorado, 
containing  0.5  oz.  Au  per  ton,  giving  70  per  cent  extraction  by  milling  and 
amalgamation,  or  of  90  per  cent  by  cyanidation.  In  comparing  the  costs 
we  have: 

SMELTING 

100  per  cent  of  0.5  oz.  Au  at $19.00        $9.50 

Mining 2 . 00 

Freight 1 . 50 

Treatment .  .  4 . 00          7 . 50 


Net  returns $2 . 00 

MILLING    AND    AMALGAMATION 

70  per  cent  of  0.5  oz.  Au  at $20 .50        $7 . 17 

Mining 2 . 00                 | 

Milling 1.00          3.00 

Net  returns $4. 17 

CYANIDATION 

90  per  cent  of  0.5  oz.  Au  at $20 .50        $9 . 23 

Mining 2 . 00 

Cyaniding 1 . 65          3 . 65 

Net  returns r .  $5 . 58 

226 


GOLD  ORE  PRICES  227 

From  the  above  comparison  it  is  seen  that  cyaniding  is  the  most  profit- 
able method  of  treatment  for  this  grade  of  ore,  and  at  this  place. 

THE  PRICE  OF  GOLD  ORE,  ALSO  COST  OF  PRODUCING  AND  SELLING 
THE  PRICE  OF  GOLD 

Price  of  Gold  Ores. — When  lead  free  or  so-called  dry  ores  containing 
gold  (and  silver)  are  sold  to  a  smelting  works  they  are  paid  for  on  the 
basis  of  dry  ore,  which  see  under  head  of  "  Purchase  of  ores,  silver-lead 
smelting." 

Costs  at  the  Belmont  Tonopah  Gold  Mill  in  1914-1915.— This  mill  of 
600  tons  daily  capacity  has,  per  ton  of  ore  put  through,  labor,  $0.419; 
supplies,  $1.318;  power,  $0.419,  using  1.68  H.P.  per  ton;  being  a  total  of 
$2.156  per  ton  of  daily  capacity. 

Costs  at  the  Homestake  Gold  Mill  in  1915. — Cost  of  stamp-milling  and 
amalgamating,  $0.2811  per  ton  of  ore.  The  tailings  from  the  mill  were 
reground  at  a  cost  in  1914  of  $0.1264  per  ton  of  product  reground.  By 
classification,  this  reground  material  yielded  two  products,  sand  and 
slime.  The  sand  was  leached  in  vats  44  ft.  diameter  at  a  cost  of  $0.1772 
per  ton.  The  other  product,  the  slime,  was  all  filter-pressed  and  all  the 
slime  plant  operating  costs  were  $0.1838  per  ton  in  1914. 

Costs  at  the  Modderfontein  Gold  Mill. — Rand  district,  South  Africa, 
m  1914,  the  cost  was  $0.686  per  ton  of  ore  milled. 

Speaking  broadly  the  treatment  cost  per  ton  of  concentrate  will  vary 
between  $2.50  and  $5  per  ton,  depending  upon  tonnage  cost  of  supplies 
delivered  at  the  plant,  also  labor.  The  extraction  of  gold,  using  the  all- 
sliming  process  varies  between  90  and  97  per  cent  on  a  raw  concentrate 
amenable  to  cyaniding,  as  at  the  Oriental  Cons.,  the  Alaska-Tread  well, 
the  Esperanza,  Waihi  and  elsewhere. 

The  price  of  gold  as  sold  to  the  mints  is  unchanging,  being  $20.67 
per  troy  ounce,  1000  fine.  From  this  the  mint  makes  a  deduction  of  2  cents 
per  ounce  to  cover  the  cost  of  melting  and  assaying. 


PART  III 
SILVER 


CHAPTER  XX 
SILVER,  ITS  ORES  AND  THEIR  TREATMENT 

Physical  Properties  of  Silver. — This  is  the  whitest  of  metals,  harder 
than  gold,  softer  than  copper,  more  malleable  and  ductile  than  any  but 
gold,  and  the  best  of  conductors  of  heat  and  electricity.  Its  specific 
gravity  is  10.5;  it  melts  at  962°  C.  and  boils  at  1850°  C.,  then  volatilizing 
and  yielding  a  green  vapor.  When  pure  and  molten  it  will  absorb  oxygen, 
which  when  the  metal  again  solidifies  causes  the  so-called  spitting  of  the 
metal,  well  known  to  assayers. 

CHARACTERISTICS  OF  SILVER  ORES 

The  silver  minerals  of  importance  in  treatment  are  as  follows : 

Native  silver,  which  sometimes  occurs  as  flakes  or  leaves,  and  as  wire- 
silver  and  metallic  silver  adherent  to  native  copper.  Native  silver  can  be 
readily  amalgamated,  but  when  present  in  particles  of  visible  size  it  is  so 
slowly  soluble  in  cyanide,  that  practically  no  extraction  can  be  obtained. 

Cerargyrite  (horn-silver,  silver  chloride),  AgCl,  is  widely  distributed. 
At  mines  it  is  found  in  the  upper  oxidized  zones?  It  is  probable  that  much 
of  the  so-called  chloride  ore  is  really  a  chjoro-bromide  (embblite).  The 
ore  is  readily  amalgamated  and  is  free-milling.  The  silver  chloride  of  it 
also  is  readily  soluble  in  cyanide  and  in  sodium  hyposulphite  solutions. 

Argentite,  Ag2&,  is  one  of  the  common  silver  ores.  By  using  chemicals 
(bluestone  and  salt)  it  can  be  amalgamated  in  pans,  and  the  silver  extracted 
thus  from  the  ore.  It  is  soluble  in  potassium  cyanide  solution. 

Stephanite,  5Ag2S,Sb2S3;  pyrargyrite,  3Ag2S,Sb2S3;  proustite, 
3Ag2S,As2Ss;  drycroasite,  AgsSb,  are  silver  sulph-arsenides  or  sulphanti- 
monides,  refractory  in  amalgamation,  even  with  chemicals,  sparingly 
soluble  in  cyanide  solution,  but  readily  soluble  in  a  solution  of  mercurous 
potassic  cyanide. 

Finally  we  have  those  silver  sulphides  that  contain  also  copper.  These 
are  polybasite,  9(Ag2Cu)S(SbAs)2Sa  and  tetrahedrite  (gray  copper  ore, 
fahlerz),  4CuFeAg2(HgZn)S,(SbAs)S3,  the  most  complex  of  all,  in  which 
the  silver  varies  from  0.06  to  31  per  cent,  being  higher  in  the  arsenical  and 
lower  in  the  antimonial  varieties.  These  sulphides  are  refractory  to  any 
amalgamation  method  and  because  of  their  copper  content  are  precluded 

231 


232  SILVER,  ITS  ORES  AND  THEIR  TREATMENT 

from  treatment  by  cyanide,  even  when  roasted.  This  does  not  interfere 
with  treatment  by  hyposulphite  lixiviation  after  roasting. 

A  number  of  rare  minerals  containing  silver  could  also  be  enumerated, 
but  for  the  metallurgist  the  minerals  above  named  are  the  important  ones. 

Silver  ores  in  general  contain  but  a  small  percentage  of  precious  metal. 
They  are  composed  mostly  of  gangue  (waste  matter  of  the  ore)  and  many 
are  treated  that  contain  less  than  0.1  to  0.2  per  cent  silver.  Thus  we  have 
at  the  Comstock  Lode,  Nevada,  silver  in  native  form  and  as  sulphide,  but 
oxides  of  iron  and  manganese  with  the  associated  sulphides,  pyrite,  blende, 
galena,  and  chalcopyrite.  At  the  Ontario  mine,  Park  City,  Utah,  the  silver 
occurs  as  argentite  and  tetrahedrite  in  a  gangue  of  quartz  and  clay  asso- 
ciated with  a  little  of  the  heavy  minerals  blende  and  galena.  These  sul- 
phides carry  silver  which  is  recovered  with  the  concentrate  in  case  of 
concentration. 

THE  EXTRACTION  OF  SILVER  FROM  ORES 

Silver  is  extracted  from  its  ores  by  milling  methods,  and  by  smelting. 
Certain  ores  contain  the  silver  in  a  form  suitable  for  cyaniding,  and  in 
consequence  that  method  of  treatment  is  coming  forward.  The  other 
methods  have  largely  dropped  out  of  use.  No  reason  appears  why 
hyposulphite  lixiviation  should  not  revive  under  the  stimulus  of  the  recent 
methods  of  agitation  and  filter-pressing.  The  patio  process  formerly 
much  practiced  in  Mexico  where  conditions  favored,  has  been  superseded 
by  cyaniding  in  many  cases,  on  account  of  the  lower  cost  of  operating  the 
latter  process,  but  in  the  past,  large  quantities  of  silver  have  been  extracted 
by  the  patio  process. 

TREATMENT  OF  SILVER  ORES 

These,  as  in  the  case  of  gold  ores,  may  be  treated  by  milling  or  by 
smelting.  Milling  processes,  with  the  exception  of  amalgamation,  are 
called  hydrometallurgical  methods.  Since  these  methods  are  often  com- 
bined with  amalgamation  and  concentration  it  would  appear  that  all 
might  better  be  grouped  under  head  of  silver  milling.  We  may  divide  the 
methods  of  silver  milling  into:  t 

A.  Amalgamation. 

(1)  Wet  silver-milling,  or  the  Washoe  process. 

(2)  Plate  and  pan  amalgamation  and  concentration. 

(3)  Dry  silver-milling ,  or  the  Reese  River  process. 

(4)  The  patio  process. 

B.  Milling  using  hydrometallurgical  processes. 

(1)  The  Augustin  process,  based  upon  the  solubility  of  silver  chloride  in 
brine. 


SILVER  ORE  TREATMENTS  233 

(2)  The  Ziervogel  process,  dependent  on  the  solubility  of  silver  sulphate 
in  hot  water. 

(3)  The  Patera  process,  in  which  silver  chloride  dissolves  in  a  solution 
of  sodium  hyposulphite. 

(4)  The  Russell  process — a  modification  of  the  Patera  process  in  which 
a  so-called  "  extra  solution  "  is  used. 

(5)  The  cyanide  process,  in  which  the  silver  minerals  either  with  or 
without  roasting,  dissolve  in  dilute  cyanide  solution. 


CHAPTER  XXI 
AMALGAMATION  OF  SILVER  ORES 

The  silver  ores  suitable  to  treat  by  milling  and  amalgamation  are  those 
that  contain  the  metal  in  such  form  as  to  be  acted  upon  by  mercury  when 
assisted  by  agitation,  heat,  and  certain  chemicals.  The  ore  is  first  crushed 
fine  by  stamps,  as  in  gold  milling,  then  treated  for  several  hours  in  grinding 
pans,  the  reactions  being  slow  compared  with  those  of  the  amalgamation 
of  gold.  In  gold  milling,  the  greater  part  of  the  gold  can  be  arrested  on 
an  apron-plate  during  the  few  seconds  in  which  the  ore  is  passing  over  it, 
while  in  silver  milling,  the  ore-pulp  has  several  hours'  contact  with 
mercury,  aided  by  heat  and  chemicals,  and  is  but  slowly  amalgamated.  In 
gold  milling,  ore  containing  0.5  oz.  Au  per  ton  can  be  profitably  milled.  In 
silver  milling,  ore  of  equivalent  value  would  contain  10  oz.  Ag  per  ton, 
or  20  times  as  much  metal.  Thus  is  seen  why  so  much  time  is  allowed  in 
silver  milling,  and  why  so  many  precautions  must  be  taken  to  be  sure  that 
all  metal  possible  is  recovered.  Several  ounces  of  silver  per  ton  often 
remain  in  the  tailing. 

The  silver  metals  suited  to  pan  amalgamation  are  cerargyrite  (horn- 
silver,  silver  chloride),  native  silver  in  flakes,  wire,  or  other  forms,  and  cer- 
tain silver  sulphides,  notably  argentite  (Ag2S).  When  the  ore  is  refractory, 
containing  arsenical  and  antimonial  sulphides,  and  especially  containing 
tetrahedrite,  galena,  or  blende,  it  is  necessary  to  roast  with  salt,  setting  free 
the  silver  or  converting  it  into  the  form  of  a  chloride,  which  becomes  sus- 
ceptible to  amalgamation.  There  is  no  sharp  line  of  demarkation  between 
free-milling  and  roasting-milling  ores.  Often  the  upper  part  of  a  vein  is 
free-milling  while  in  depth  base  metals  and  sulphides  begin  to  come  in, 
and  it  finally  becomes  necessary  to  roast  the  ore.  The  best  extraction 
therefore  is  obtained  from  decomposed  or  oxidized  ore,  in  which  trie  silver 
materials  occur  in  a  form  that  renders  possible  the  action  of  the  mercury. 
There  are  few  deposits  of  oxidized  ores  containing  silver  chloride  and  native 
silver  that  as  a  whole  are  suitable  for  free  silver  milling.  Such  ore,  so  far 
as  silver  chloride  is  concerned,  can  also  be  treated  by  cyanidation,  but  the 
latter  method  would  not  recover  native  silver. 

Arsenic  and  antimony  compounds  interfere  with  amalgamation  by 
fouling  the  quicksilver,  checking  the  reactions  of  the  chemicals  added  to 

234 


WET-SILVER  MILLING  235 

promote  amalgamation,  and  by  carrying  off  silver,  which  is  incapable  of 
being  amalgamated  with  them. 

The  Washoe  process,  developed  through  the  combination  of  the  Cali- 
fornia stamp  mill  with  an  elaboration  of  the  Norwegian  Tina  for  fine 
grinding  and  amalgamation  (pan)  and  the  chemicals  of  the  Patio  process, 
was  introduced  for  the  treatment  of  Comstock  ores  in  1860.  This  amal- 
gamation process,  adapted  to  American  conditions,  rapidly  assumed  the 
same  position  in  the  United  States  for  the  treatment  of  silver-gold  ores 
as  was  occupied  by  the  patio  process  in  Mexico  for  the  treatment  of  the 
same  class  of  ores. 

A  later  development  when  treating  complex  ores  was  to  give  a  chlorid- 
izing  roast  preceding  grinding  and  amalgamation  in  the  pans.  This 
was  known  as  the  Reese  River  process. 

(1)     WET  SILVER-MILLING  WITH  TANK-SETTLING 

This  is  also  known  as  the  Washoe  process,  receiving  the  name  from  the 
place  where  it  was  perfected  for  the  treatment  of  ores  from  the  Comstock 
Lode,  Nevada.  The  process  is  applicable  to  the  so-called  free-milling  ores, 
in  which  the  silver  occurs  native,  as  chloride  or  in  small  amount  as  argen- 
tite.  The  ore  should  be  free  from  lead  and  from  any  tough  clayey  gangue. 

In  wet  silver-milling,  the  process  consists  in  coarse-crushing  the  ore, 
stamping  it  fine,  and  collecting  it  in  settling-tanks.  The  crushed  sand  is 
ground  in  amalgamating-pans  using  mercury  to  collect  the  silver.  The 
sand  is  separated  from  the  silver-bearing  mercury  in  settling-pans  and  is 
rejected.  The  amalgam  is  strained  from  the  mercury,  retorted,  and  the 
retort-residue  melted  into  silver  ingots.  Gold  present  in  the  ore  is  recov- 
ered as  well  as  the  silver.  The  process  resembles  gold-milling  except  that 
amalgamation  and  the  removal  of  the  amalgam  is  effected  in  pans. 

Fig.  134  is  a  sectional  elevation  of  a  wet-crushing  tank-mill  for  the 
treatment  of  free-milling  silver  ores.  The  ore  from  the  mine  is  amal- 
gamated on  plates  precisely  as  in  gold  milling,  which  see.  The  coarse  crush- 
ing is  done  during  the  ten-hour  day-shift. 

Water  (6  to  8  tons  per  ton  of  ore)  is  at  the  same  time  supplied  in  the 
mortar,  and  forms  a  pulp,  which  is  splashed  through  the  30-mesh  screens 
by  the  motion  of  the  stamps  using  a  double-discharge  mortar.  The  large 
screen-opening  possible  with  a  double  discharge  favors  a  more  rapid  pul- 
verization than  would  be  possible  with  a  single-discharge  mortar.  The  pulp 
flows  by  launders  /  into  the  settling-boxes  or  tanks  g  7  ft.  square  by  3  ft. 
deep.  There  is  a  double  row  of  these  tanks,  twenty  hi  a  row,  occupying 
the  length  of  the  mill  in  front  of  the  stamps.  The  flow  of  the  pulp  is 
from  box  to  box  in  series,  until  it  goes  by  launder  to  a  settling-pond  out- 
side the  mill.  Most  of  the  solids  settle  in  the  first  boxes,  a  further  portion 


236 


AMALGAMATION  OF  SILVER  ORES 


WET  SILVER  MILLING 


237 


dropping  in  the  succeeding  ones,  and  the  turbid  water  passing  to  the  pond. 
Here  it  has  its  final  chance  to  settle  before  running  to  waste,  or  it  may  be 
again  used  in  the  mill  if  water  is  so  scarce  that  it  pays  to  do  this.  The 
settled  slime  is  dug  from  the  pond  at  a  later  time  and  treated  like  the  rest 
of  the  crushed  ore. 

A  variation  of  this  method,  shown  in  Fig.  134,  consists  in  conducting 
the  flow  from  the  last  box  gr  by  an  inclined  elevator  to  a  tank  h  situated 
in  front  of  and  above  the  battery,  the  dirty  water  being  again  used  for 
stamping.  When  the  first  settling-box  is  full  the  flow  of  the  pulp  is  by- 
passed into  the  next  one.  The  contents  of  the  full  box  are  shoveled  upon  the 
floor  adjoining,  and  thence  taken  as  needed  to  the  amalgamating-pans  q. 
The  emptied  box  has  the  flow  of  the  last  one  turned  into  it,  thus  making 
it  the  last  in  the  series,  and  the  launders  are  so  arranged  that  this  can  be 
done. 

The  ore,  thrown  out  upon  the  floor,  is  fed  directly  into  the  pans  or  loaded 
into  the  tram-car  seen  in  Fig.  134  and  conveyed  to  them. 

Fig.  135  represents  a  pan. 
It  is  5  ft.  diameter  by  30  in. 
deep,  and  is  furnished  with  a 
central  sleeve  or  cone  through 
which  rises  a  shaft  carrying  a 
cylindrical  casting  called  a 
spider,  which  becomes  bell- 
shaped  and  broadens  into  feet 
below.  The  spider  carries, 
bolted  to  the  feet,  a  flat  cast- 
iron  ring  called  a  muller,  and 
to  the  under  side  of  the  muller 
are  attached  six  shoes  or  plates 
of  chilled  cast-iron  2J  in.  thick. 
The  spider,  muller,  and  shoes 
are  raised  or  lowered  as  de- 
sired, by  means  of  a  hand- 
wheel  and  screw  at  the  top  of 
the  shaft,  which  is  driven  by 
bevel  gearing  from  the  horizontal  shaft  and  pulley  below.  Upon  the 
bottom  of  the  pan  rest  chilled  cast-iron  plates  or  dies  that  furnish  the 
lower  or  fixed  grinding  surface.  The  shoes  attached  to  the  muller  revolve 
60  R.P.M.  and  rubbing  upon  the  dies,  grind  the  ore. 

In  working  the  pans,  the  shoes  are  raised  J  in.  from  the  dies  and  set  in 
motion,  the  pan  is  partly  filled  with  water,  and  3000  Ib.  of  the  damp  pul- 
verized ore  is  shoveled  in.  The  ore  and  water  nearly  fill  the  pan  and  the 
mixture  is  stirred  until  it  is  of  the  consistence  of  honey.  The  motion  estab- 


FIG.  135. — Five-foot  Continuous  Grinding  Pan. 


238  AMALGAMATION  OF  SILVER  ORES 

lishes  a  movement  or  current  of  pulp  beneath  the  muller  toward  the  periph- 
ery. At  the  periphery  it  rises,  flows  toward  the  center,  sinks,  and  passes 
again  under  the  shoes.  To  assist  the  action,  the  rising  pulp  is  deflected 
inward  by  cast-iron  wing-plates. 

After  thorough  mixing  in  the  pan  the  shoes  are  lowered  until  they 
touch  the  dies,  and  grinding  goes  on  for  1|  hours,  the  content  of  the  pan 
being  meanwhile  heated  nearly  to  boiling  by  steam  under  pressure  from  a 
pipe  that  dips  beneath  the  surface  of  the  charge,  the  pan  being  covered. 

After  grinding,  the  shoes  are  raised  and  300  Ib.  of  mercury  (10  per  cent 
the  weight  of  the  ore)  is  added,  by  sprinkling  it  through  a  fine  strainer. 
The  mixing  is  then  continued  four  hours.  The  mercury  takes  up  silver 
most  rapidly  at  first,  but  the  action  afterward  slackens.  The  globules  of 
mercury  suspended  in  the  pulp  take  the  silver  as  they  come  in  contact 
with  it.  Care  is  taken  to  have  the  pulp  of  the  right  consistence  so  that 
mercury  will  not  settle  out.  This  condition  is  shown  when  a  wooden 
stick,  dipped  in  the  pulp  and  withdrawn,  is  found  to  be  covered  with  a 
thick  mud  in  which  are  disseminated  minute  globules  of  mercury.  If  the 
ore  is  refractory,  salt  and  copper  sulphate  are  advantageously  added  at  the 
beginning  of  grinding  to  accelerate  the  reactions,  promote  amalgamation, 
and  increase  the  yield  of  silver. 

The  charge  above  treated  having  been  amalgamated,  the  pan  is  ready 
to  empty  into  the  settler  r.  About  fifteen  minutes  before  the  discharging, 
the  speed  of  the  muller  is  reduced  to  40  R.P.M.  and  the  pan  filled  to  the 
top  with  water.  A  plug  closing  the  discharge  opening  at  the  bottom  of  the 
pan,  seen  at  the  left  in  section,  Fig.  135,  is  pulled  out,  and  the  entire  content 
run  by  launder  to  an  8-ft.  settler,  at  a  lower  level,  shown  at  the  right  of 
the  amalgamating-pans  in  Fig.  134.  Emptying  the  pan  and  washing  it  with 
a  hose  takes  half  an  hour,  after  which  time  the  plug  is  replaced,  and  the 
pan  is  ready  for  another  charge.  Thus  the  total  time  for  the  cycle  of  opera- 
tions described  is  six  hours,  making  it  possible  to  treat  four  charges  daily. 

The  reactions  that  take  place  in  the  pan  are  as  follows : 

Native  silver  in  threads,  films,  flakes,  or  grains  readily  combines  with 
the  mercury  and  forms  an  amalgam  which  contains  a  large  excess  of  mer- 
cury. 

Silver  chloride  in  contact  with  the  mercury  decomposes  as  follows : 

(1)  2AgCl+2Hg  =  Hg2Cl2+2Ag. 

The  metallic  silver  liberated  amalgamates  with  additional  mercury. 
The  particles  of  iron,  abraided  from  the  stamps  and  the  bottom  of  the  pan, 
decompose  the  mercury  salt  and  liberate  the  mercury  as  follows: 

(2)  Hg2Cl2+Fe  =  FeCl2+2Hg. 


WET  SILVER  MILLING  239 

Many  so-called  free-milling  silver  ores  contain   argentite  which  in  part 
is  decomposed  by  mercury  as  follows: 


(3)  Ag2S+2Hg  =  Ag2+Hg2S. 

The  sulphide  of  mercury  thus  formed  is  lost.  We  have  already  stated  that 
chemicals,  notably  copper  sulphate  and  common  salt,  are  added  to  promote 
the  decomposition  of  the  silver  sulphide.  There  is  added  in  the  amal- 
gamating-pan  from  6  to  18  Ib.  salt  and  from  3  to  9  Ib.  copper  sulphate  per 
ton  of  ore  treated.  The  reactions  as  generally  given  are  the  following: 

(4)  CuSO4+2NaCl  =  Na2SO4+CuCl2. 

The  chloride  of  copper  acting  on  the  silver  sulphide  decomposes  it: 

(5)  Ag2S+CuCl2  =  CuS+2AgCl. 

The  silver  chloride  amalgamates  as  shown  by  reaction  (1). 

The  complete  separation  of  the  mercury  with  the  silver-amalgam  is 
effected  in  the  settler,  there  being  one  settler  provided  for  two  amalga- 
mating-pans.  The  settler  is  8  ft.  diameter  by  3  ft.  deep,  three  times  the 
capacity  of  the  amalgamating  pan,  but  of  similar  construction,  as  shown 
in  Fig.  136.  No  grinding  is  required,  but  the  pulp  must  be  agitated  with  the 
wooden  shoes  with  which  the  settler  is  provided.  The  shoes  nearly  touch 
the  bottom  of  the  settler,  the  exact  height  being  adjustable.  The  grooved 
border  at  the  bottom  just  within  the  sides  of  the  settler  has  a  slight  grade 
to  the  outlet  and  mercury-well  at  the  left.  The  mercury  settles  from  the 
pulp,  flows  to  the  lowest  point  and  stands  at  a  height  that  balances  the 
hydrostatic  head  of  the  content  of  the  pan.  Since  the  specific  gravity  of 
mercury  is  14  and  the  content  of  the  settler  approximately  1.5,  the  height 
of  the  mercury  is  a  little  less  than  4  in.  The  bottom  outlet-hole  of  the  well 
in  plugged.  At  different  heights  in  the  side  of  the  pan  there  are  provided 
openings  that  are  kept  closed  by  plugs.  When  the  plugs  are  withdrawn 
the  tailing  and  water,  free  from  mercury,  pass  out  of  the  pan. 

The  shoes  of  the  settler  having  been  set  in  motion,  at  the  rate  of  15 
R.P.M.,  and  raised  8  in.  above  the  bottom,  the  contents  of  the  two  pans 
are  run  in,  as  has  been  described.  Water  is  then  added  to  within  6  in.  of 
the  top,  greatly  thinning  the  pulp,  and  filling  the  settler.  After  half  an 
hour  the  shoes  are  gradually  lowered  until,  at  the  end  of  two  hours,  they 
nearly  touch  the  bottom.  The  purpose  of  the  agitation  is  to  keep  the 
lighter  portion  of  the  ore  (now  called  the  tailing)  in  suspension,  while  the 
silver-bearing  mercury,  the  heavier  particles  of  sulphide,  and  the  particles 
of  iron  from  the  stamps  collect  at  the  bottom.  The  stirring  is  continued 
3J  hours,  after  which  the  highest  plug  in  the  side  of  the  settler  is  removed, 
and  the  turbid  water  containing  tailing  is  allowed  to  escape  by  launder,  a 
stream  of  clear  water  being  meanwhile  allowed  to  flow  through.  The 


240 


AMALGAMATION  OF  SILVER  ORES 


plugs  are  then  withdrawn  one  by  one  until  the  settler  is  emptied  of  all  the 
content  except  the  heavy  portion  containing  sulphide,  iron  particles,  and 
the  mercury.  Emptying  takes  half  an  hour,  and  the  cycle  of  operations 
becomes  six  hours  as  in  the  case  of  the  amalgamating  pan.  Since  escaping 
tailing  contains  sulphide,  it  may  be  run  over  riffles,  or  blanket-lined  laun- 
ders, before  running  to  waste. 

The  silver-bearing  mercury  or  diluted  amalgam,  a  mixture  of  silver- 
amalgam  and  mercury,  collecting  in  the  mercury  well,  overflows  by  an 


FIG.  136. — Eight-foot  Settler. 

escape-opening  indicated  in  Fig.  136.  From  the  opening  it  passes  by  a  half- 
inch  pipe  to  the  amalgam  safe  shown  at  the  right  of  the  settler,  Fig.  134. 
The  safe,  arranged  to  prevent  theft  of  the  amalgam,  is  shown  on  a  larger 
scale  in  Fig.  137.  The  amalgam  and  mercury  'enter  a  conical  canvas  sack 
or  filter.  The  mercury  oozes  through  the  pores  of  the  canvas  while  the 
amalgam  containing  as  little  as  14  per  cent  silver  is  retained.  Occasionally, 
after  amalgam  has  accumulated,  the  sack  is  squeezed  between  the  hands  to 
remove  the  surplus  mercury,  and  the  compressed  amalgam  containing  20 


TREATMENT  OF  SILVER  AMALGAM 


241 


to  28  per  cent  silver,  is  reserved  for  retorting.  The  mercury  flows  out  at  the 
bottom  through  an  outlet  provided,  as  seen  in  Fig.  137,  and  is  collected  at  a 
lower  level  in  the  boot  w,  Fig.  134,  of  the  mercury  elevator,  shown  at  the 
right.  The  elevator  discharges  to  a  mer- 
cury tank  s  commanding  the  amalgamat- 
ing pans,  to  which  it  is  delivered  as  needed 
through  the  pipe  shown  in  the  figure.  Over 
the  stamps  and  the  pans  are  seen  the  over- 
head tracks  that  carry  crawls  by  which 
the  heavy  parts  of  the  machines  are  lifted 
or  transferred.  This  facilitates  the  work  of 
repairs  and  replacements. 

The  loss  of  mercury  is  commonly  1  to 
1.5  Ib.  per  ton  of  ore  treated.  A  part  is 
lost  in  handling,  but  the  principal  loss  is 
the  flouring,  which  causes  the  mercury  to 
escape  in  the  tailing.  'The  loss  is  greater 
with  talcose  or  clayey  ores,  and  in  those 
carrying  cerussite,  chalcopyrite,  or  galena. 
Loss  is  caused  by  grease  coating  the  par- 
ticles of  mercury,  in  case  this  enters  the 
ore  from  the  machinery. 

Treatment  of  the  Amalgam. — Since  the 


FIG.  137.— Amalgam  Safe. 


weight  of  metal  recovered  in  silver  milling  is  much  greater  than  in  gold 
milling,  the  retorting  of  amalgam  must  be  performed  on  a  larger  scale. 
Fig.  138  shows  a  sectional  elevation  and  a  plan  of  a  combined  retorting  and 
melting  furnace  with  the  overhead  crawl  and  chain-blocks  by  which  the 
large  melting  crucibles  are  lifted  from  the  fire  in  the  melting  furnace  and 
transferred  for  pouring.  At  the  left  is  shown  in  the  elevation  a  cross- 
section  of  the  cast-iron  cylindrical  retort  which  is  10  in.  diameter  inside 
by  28  in.  long,  resting  upon  arched  cast-iron  supports.  There  is  a  horizon- 
tal pipe,  and  a  vertical  water-cooled  pipe,  not  shown  in  the  illustration,  in 
which  the  mercury  condenses  and  from  which  it  falls  into  a  tub  of 
water  below.  As  seen  hi  the  plan,  the  front  end  of  the  retort  is  provided 
with  a  cover  which  can  be  securely  clamped  in  position. 

The  charge  of  amalgam,  containing  20  per  cent  mercury,  should  weigh 
500  Ib.  and  only  half  fill  the  retort.  After  filling,  the  cover  is  clamped  on, 
first  luting  the  joint  with  flour  paste.  A  wood  fire  is  started  on  the  grate 
under  the  retort.  The  temperature  is  kept  low  at  first,  increasing  to  a  red- 
heat  at  the  end,  f  to  J  cord  of  wood  being  used.  The  operation  lasts  ten 
to  fourteen  hours,  care  being  taken  not  to  heat  the  retort  rapidly,  nor, 
for  fear  of  blistering  it,  to  raise  the  temperature  too  high.  The  fire  is 
then  allowed  to  burn  down,  and  the  retort  to  cool.  The  lid  is  taken  off 


242 


AMALGAMATION  OF  SILVER  ORES 


and  the  silver  residue  removed  to  a  crucible.  This  is  seized  by  basket- 
tongs,  which  clasp  it  firmly  so  that  it  can  be  lifted  by  the  chain-hoist, 
transferred  by  the  crawl  to  the  ingot  mold,  and  poured.  These  molds, 
11  in.  long  by  4J  in.  wide  and  deep,  hold  1000  oz.,  or  70  Ib.  silver. 


FIG.  138. — Horizontal  Retort  and  Melting  Furnace  for  Silver  Mill. 

The  settler  tailing  contains  heavy  unaltered  ore  which  may  be  con- 
centrated to  recover  heavy  sulphides  and  particles  of  amalgam. 

Costs. — The  costs  of  pan-amalgamation  with  tank-settling  (Washoe 
process)  in  1910  per  ton  of  ore  treated  is: 


THE  BOSS  PROCESS  243 

Power $0.087 

Labor 0.361 

Chemicals  (salt,  acid,  Milestone) 0 . 465 

Loss  of  mercury .\ 0 . 750 

Wear  of  pans 0.200 

Wear  of  dies  and  shoes 0 . 400 

Oil,  interest,  and  superintendence 0 . 100 


Total  cost  per  ton. .  .   $2 . 363 

One  notes  in  particular  the  larger  cost  of  supplies  (chemicals,  mercury, 
and  castings)  compared  with  like  items  in  gold  milling. 

THE  BOSS  PROCESS  OF  SILVER  MILLING 

This  system,  originated  by  M.  P.  Boss,  a  California  engineer, 
differs  from  the  Washoe  process  in  being  continuous  and  generally 
requiring  less  labor.  However,  the  Allis-Chalmers  Co.  has  designed, 
for  the  Washoe  process,  a  wet-crushing  mill  in  which  the  settling- 
boxes  have  sloping  bottoms,  so  arranged  that  the  content  is  trans- 
ferred to  the  pans  with  but  little  labor.  This  takes  away  the  advan- 
tage urged  in  favor  of  the  Boss  system.  It  may  be  added  that  the  settling 
of  the  pulp  in  large  tanks,  combined  with  a  mechanical  system  of  excavating 
the  content  as  hi  the  cyanide  process,  ought  to  be  efficient  and  labor-saving. 
The  Boss  system  may  be  applied  to  free  milling  ores  and  to  refractory  ores 
that  need  to  be  first  roasted. 

THE  HIGH-GRADE  NIPISSING  MILL,  COBALT,  ONTARIO 

The  ore,  containing  native  silver  and  argentite,  together  with  the  arsen- 
ides of  cobalt  and  nickel  (6  per  cent  Ni,  7  to  8  per  cent  Co,  40  per  cent  As), 
after  being  crushed  to  70-mesh  at  the  sampling  mil],  is  delivered  at  the  plant 
with  an  average  content  of  2600  oz.  silver  per  ton.  It  is  fed  to  a  tube-mill 
(see  Fig.  149),  20  ft.  long  by  4  ft.  diameter.  The  charge  consists  of  3J 
tons  of  ore,  4J  tons  of  mercury,  and  a  5  per  cent  cyanide  solution.  The 
tube-mill  is  closed  at  both  ends.  Air,  to  accelerate  chemical  action,  is  intro- 
duced through  a  pipe.1  There  is  also  an  ingenious  device  whereby  the 
excess  of  air  is  subsequently  expelled.  After  nine  hours  in  the  tube-mill,  98 
per  cent  of  the  silver  has  been  extracted  from  the  ore,  which,  in  the  form 
of  pulp,  then  passes  to  a  settler,  where  the  amalgam  is  separated  by 
gravity.  Thence  it  goes  to  a  clean-up  pan  and  drainers.  These  last  are 
canvas  bags  for  removing  any  excess  of  mercury. 

The  pulp  and  solution,  deprived  of  amalgam,  pass  to  a  vat  and  are 
fed  to  a  Butters  filter,  the  clarified  solution  going  to  zinc  boxes  where 

1  As  the  result  of  the  oxidation  of  the  arsenides  the  temperature  of  the  charge  would 
rise  to  the  boiling-point  were  not  the  air  supply  to  the  barrel  controlled. 


244  AMALGAMATION  OF  SILVER  ORES 

the  dissolved  silver  is  precipitated  on  zinc  shavings,  thus  obtaining  an 
additional  2  per  cent  recovery.  The  shavings  are  in  the  form  of  coarse 
wire,  necessary  on  account  of  the  strength  of  the  cyanide  solution.  The 
residue,  left  on  the  filter,  containing  8  to  9  per  cent  cobalt,  is  afterwards 
sold  for  the  value  of  this  metal  plus  85  per  cent  of  the  silver  contents, 
so  that  from  ore  of  2600  oz.  silver  per  ton,  only  about  4  oz.  of  silver 
value  is  lost. 

The  amalgam,  containing  80  per  cent  mercury  and  20  per  cent  silver, 
is  placed  in  retorts,  each  of  which  holds  450  Ib.  After  the  mercury  has 
been  distilled,  the  silver,  still  containing  1  per  cent  mercury,  is  taken  to  a 
reverberatory  furnace.  Here  it  is  melted  in  a  charge  of  25,000  oz.  After 
fifteen  hours'  exposure  to  a  hot  oxidizing  atmosphere,  without  addition  of 
any  flux,  the  molten  metal  is  cast  in  ingots,  each  weighing  1100  oz.  silver, 
which  is  999  fine.  Two  oil-burners  afford  the  necessary  heat.  The  flue 
from  the  furnace  is  provided  with  a  water-jet  condenser,  whereby  1000  to 
2000  Ib.  mercury  is  arrested  monthly.  The  gases  escape  at  100°  F.  Dur- 
ing February,  1912,  550,000  oz.  of  silver  was  melted  in  this  small  plant. 

The  richness  of  the  mine  product  under  treatment  and  the  complete- 
ness of  the  metallurgical  operations  leave  a  vivid  impression.  Within  a 
small  building  it  was  possible  to  watch  the  successive  stages  by  which  a 
complex  ore  of  a  refractory  type  yielded  its  precious  content  in  metal  of 
such  purity  as  to  be  ready  for  the  mint.  The  entire  process  is  so  expe- 
ditious that  the  silver  is  delivered  at  New  York  within  a  week  of  the 
day  when  the  ore  is  received  at  the  mill  and  payment  for  the  yield  is 
received  concurrently  with  the  shipment.  No  less  than  20  tons  of  mer- 
cury is  in  use  at  a  given  time.  The  cyanide  has  a  cleansing  action  upon 
it ;  indeed,  the  use  of  mercury  would  be  impracticable  without  the  cyanide, 
for  the  mercury  would  become  "  sick  "  or  fouled,  so  as  to  hinder  amalga- 
mation with  the  silver  in  the  finely  ground  arsenical  ore.  The  yoking  of 
cyanidation  and  amalgamation  constitutes  another  remarkable  feature. 

The  right  half  of  the  flow-sheet,  Fig.  149,  gives  a  clear  idea  of  the 
progress  of  operations.  The  product  received  consists  of  two-thirds 
hand-picked  ore  of  2800  oz.  per  ton,  the  rest,  jig  products  of  an  average 
value  of  2400  oz.  per  ton. 

t 

(2)     PLATE  AND  PAN— AMALGAMATION  AND  CONCENTRATION   OF 

SILVER  ORES 

This  is  used  on  ores  carrying  silver,  gold,  and  sulphides  of  the 
heavy  metals,  such  as  galena,  blende,  and  pyrite,  and  sulphides  which 
contain  silver  and  gold.  It  is  necessary  that  the  silver  not  in  the  sulphides 
be  amalgamable,  as  is  silver  chloride,  argentite  or  native  silver, 
i  The  process  consists  in  wet-stamping  the  ore,  running  the  pulp  over 
apron-plates  as  in  gold  milling,  concentrating  the  sulphides,  which  are 


AMALGAMATION  AND  CONCENTRATION 


245 


shipped  to  the  smelter,  and,  as  in  the  Washoe  process,  pan-amalgamating 
the  tailing  and  retorting  the  amalgam  to  recover  the  silver  and  gold. 

Compared  with  either  wet  or  dry  silver-milling,  the  process  has  much 
to  commend  it.  The  ore  being  refractory,  the  wet  process  would  recover 
little  value.  The  tonnage  stamped  by  the  dry  method  with  roasting  would 
be  low  compared  with  wet-stamping,  which  is  one  and  one-half  to  twice  as 
rapid.  It  is  true  that  by  dry-stamping  and  roasting  we  are  able  to  extract 
at  least  10  per  cent  more  metal  than  can  be  obtained  by  raw  amalgamation, 
but  this  is  offset  by  the  cost  of  treatment  and  the  loss  of  precious  metal  in 
roasting.  The  process  also  saves  lead  and  removes  galena,  sulph-arsenides, 


FIG.  139. — Stamp  Mill  Using  Amalgamation  and  Concentration. 

and  sulph-antimonides,  all  of  which  tend  to  foul  and  cause  the  loss  of  mer- 
cury. Such  minerals  are  not  amenable  to  amalgamation,  and  by  removing 
them  for  smelting  there  results  a  cleaner  or  higher-grade  bullion.  Man- 
ganese minerals  that  consume  chemicals  in  the  pan  are  also  removed  by 
concentration. 

Amalgamation  and  Concentrating  Mill. — Fig.  139  is  a  perspective  view 
of  a  10-stamp  mill.  Let  us  suppose  we  are  to  treat  an  ore,  in  part  oxidized, 
but  containing  the  heavy  minerals  of  lead  and  copper,  with  pyrite,  arsenides, 
and  manganese  minerals.  The  ore  contains  the  precious  metals,  a  gangue 
of  quartz,  calcite,  and  a  little  clay,  and  disseminated  through  it  gold  and 
the  amalgamable  silver  minerals  cerargyrite,  argentite,  and  native  silver. 


246  AMALGAMATION  OF  SILVER  ORES 

The  purpose  is  to  save  the  precious  metals  by  plate  and  pan-amalgamation, 
and  the  heavy  minerals  with  silver  and  gold  by  concentration.  Some  of 
the  silver  and  gold  escapes  recovery  and  is  lost  in  the  tailing.  Since  sulph- 
arsenides  and  manganese  minerals  are  mostly  removed,  they  do  not  inter- 
fere with  subsequent  pan-amalgamation  where  arsenic  would  sicken  the 
mercury  and  manganese  consume  chemicals. 

The  ore  and  water  are  fed  automatically  to  a  10-stamp  battery,  each 
stamp  crushing  4  tons  per  twenty-four  hours  to  pass  a  30-mesh  screen. 
The  pulp  issuing  from  the  mortar  flows  over  two  apron-plates  (one  for  each 
five-stamp  mortar)  and  a  part  of  the  gold  and  silver  is  recovered.  The 
flow  is  distributed  evenly  to  four  concentrating  tables  at  a  lower  level,  the 
concentrate  (10  per  cent  of  the  whole)  being  separated  to  ship  to  smelting 
works,  while  the  tailing  is  carried  to  the  ten  settling-boxes  in  a  double  row. 
These  are  seen  at  the  left  of  the  pans.  The  distribution  is  into  a  double 
launder  between  the  two  rows.  By  drawing  the  plugs  in  the  bottom  of 
the  launder,  the  flow  can  be  directed  into  any  box  desired.  From  this 
point  on,  the  operation  is  conducted  as  described  for  the  Washoe  process. 
There  are  four  amalgamating-pans  and  two  settlers.  Bluestone  and  salt 
are  used  to  decompose  the  argentite.  Mercury  or  amalgam  escaping  the 
apron-plates  finds  its  way  into  the  settling-boxes  and  thence  to  the  pans, 
and  is  more  thoroughly  recovered  than  if  it  depended  upon  obtaining  it 
in  the  concentrate  as  hi  gold-milling.  The  four  5-ft.  combination  grind- 
ing and  amalgamation-pans  each  treat  3000  Ib.  per  charge,  and  with  a 
four-hour  treatment,  this  equals  36  tons  daily,  which  with  the  4  tons  of 
concentrate  already  mentioned  is  a  40-ton  output  of  the  mill.  Some  ores, 
not  so  readily  treated,  take  six  to  eight  hours,  and  lessen  the  capacity  of 
the  mill  accordingly. 

In  a  certain  ore  of  this  kind,  containing  0.40  oz.  Au  and  9.02  Ag  per  ton 
the  recoveries  on  the  apron  plates  were  22  per  cent  of  the  gold  and  3  per 
cent  of  the  silver  respectively;  at  the  concentrating  tables  28  and  32  per 
cent;  in  the  pans  and  32  and  35  per  cent;  lost  in  the  tailing  18  and  30 
per  cent.  Of  the  lead  and  copper  85  per  cent  was  saved  in  the  concentrate. 

Concentrating  adds  but  little  to  the  cost  of  this  milling,  so  that  $3 
per  ton  may  be  taken  as  a  fair  estimate  in  1910. 

I 

THE  CHLORIDIZING  ROASTING  OF  SILVER  ORES 

Silver  ore  containing  sulph-arsenides,  sulph-antimonides,  or  tetrahe- 
drite,  cannot  be  treated  directly  by  amalgamation  nor  by  cyaniding. 
Such  ore  is  subjected  to  a  roast  with  salt  to  convert  the  silver  into  a  chloride, 
before  it  can  be  successfully  treated  by  these  methods.  The  above  min- 
erals are  often  accompanied  by  pyrite,  blende,  chalcopyrite,  and  galena. 

Preliminary  to  roasting,  such  ore  is  dry-crushed,  either  by  rolls  or  by 


CHLORIDIZING  ROASTING  247 

stamps.  Ores  containing  galena  and  blende  are  preferably  crushed  to 
40-mesh  size,  those  having  pyrite  to  8  to  10  mesh.  The  "  Roasting  "  is  done 
in  a  reverberatory  furnace,  and  requires  the  use  of  salt.  There  must 
also  be  3  to  8  per  cent  pyrite  present  to  furnish  sulphur  for  the  reactions, 
and  if  the  ore  does  not  contain  this,  it  must  be  added.  If  more  than  8 
per  cent  sulphur  is  present,  the  percentage  is  reduced  to  that  point  by 
roasting  before  the  salt  is  added.  The  amount  of  salt  required  varies 
according  to  the  quantity  of  copper  and  iron  sulphides  present  which  con- 
sume the  evolving  chlorine. 

Chloridizing  Roasting. — This  operation  is  at  first  an  oxidizing  one  con- 
ducted at  the  temperatures  specified  in  the  chapter  on  roasting.  The 
action  is  chiefly  upon  the  heavy  metals,  converting  them  into  either  oxides 
or  sulphates.  It  may  be  divided  into  three  stages:  (1)  the  kindling,  (2) 
the  desulphurization,  and  (3)  the  chlorination  of  the  ore. 

First  Stage. — In  the  first  or  kindling  stage  we  find  the  loosely  held 
sulphur  being  driven  off,  and  the  ore  taking  fire,  producing  a  blue  flame. 

Second  Stage. — In  the  second  stage,  the  air  oxidizes  the  sulphides,  and 
particularly  the  newly  formed  iron  sulphide.  Reacting  upon  the  sulph- 
antimonides  and  arsenides,  it  volatilizes  them  and  removes  them  from  the 
ore.  Copper  and  iron  sulphates  are  also  formed,  the  latter  according  to 
the  following  reaction ; 

(6)  3FeS+ HO  =  2SO2+Fe2O3+FeSO4. 

Third  Stage. — In  the  third  stage,  at  590°  C.,  the  sulphate  formed  in 
conjunction  with  air  reacts  upon  the  salt,  thus: 

(7)  FeSO4+2NaCl  =  N2SO4+FeCl2 
and 

(8)  4FeCl2+3O  =  2Fe2O3+4Cl2. 

The  chlorine  thus  liberated  acts  at  once  upon  the  silver  compounds 
and  converts  them  into  chlorides.  Thus: 

(9)  Ag2S+02+Cl2  =  2AgCl+S02. 

Zinc  blende  becomes  oxide  and  zinc  sulphate,  while  sulphur  dioxide 
escapes.  Galena  and  zinc  sulphate  remain  inactive  and  fail  to  decompose 
the  salt.  They  roast  slowly,  while  pyrite,  in  presence  of  salt,  decomposes 
quickly,  and  generates  chlorine  at  a  period  in  the  roasting  when  neither  the 
blende  nor  galena  is  sufficiently  oxidized  to  expose  silver  to  the  action  of  the 
chlorine.  If,  therefore,  the  salt  is  mixed  with  the  ore  at  the  battery,  the 
chlorine  generated  by  the  reaction  of  the  ferrous  sulphate  and  salt  is  lost, 
an  imperfect  chlorination  results,  no  matter  how  long  roasting  is  con- 
tinued, nor  how  much  salt  is  added.  Hence  in  roasting  an  ore  containing 


248  AMALGAMATION  OF  SILVER  ORES 

blende  and  galena  it  is  of  the  greatest  importance  to  add  the  salt  later  and 
not  at  the  battery.  On  the  other  hand,  if  the  roasting  continues  until 
the  sulphides  are  well  oxidized,  the  iron  sulphate  decomposes  and  no 
chlorine  is  generated,  and  again  we  have  a  badly  chloridized  ore.  The 
desirable  time  to  add  the  salt  is  after  continued  roasting  at  a  low  heat  that 
does  not  break  up  the  iron  sulphate.  This  is  shown  when  the  black  color 
of  the  ore  changes  to  brown,  but  shows  still  the  presence  of  black  particles. 
A  distinct  odor  of  chlorine  is  then  to  be  noticed,  due  to  the  decomposition 
of  the  salt.  The  best  results  could  be  obtained  by  adding  a  mixture  of 
green  vitriol  (ferrous  sulphate)  and  salt;  but  the  ore  would  hardly  justify 
the  expense. 

The  salt  is  added  to  the  dry  ore  at  the  time  of  charging,  if  the  percentage 
of  sulphur  is  suitable,  or  later  if  the  excess  of  sulphur  must  be  first  removed 
by  roasting.  The  temperature  is  increased  only  gradually  to  kindle  or 
start  the  ore  to  burning  and  to  begin  oxidation.  As  the  temperature  rises 
oxidation  and  the  formation  of  sulphates  occur,  and  at  the  necessary  high 
temperature  these  act  upon  and  decompose  the  salt  and  chloridize  the  ore. 

Heap  Chlorination. — It  is  not  considered  necessary  to  continue  the 
roasting  to  convert  all  possible  silver  into  chloride,  but  to  withdraw  the 
charge  while  hot  before  this  stage  is  reached.  During  the  gradual  cooling 
(twelve  to  thirty  hours)  further  chloridizing  proceeds,  due  to  the  mass 
action  of  the  free  chlorine,  with  which  the  ore  is  saturated,  acting  on  the 
undecomposed  silver  sulphide.  This  may  increase  the  chloridization  10  to 
40  per  cent. 

Upon  completion  of  the  operation  of  "  heap  chlorination,"  as  it  is 
called,  and  with  ores  containing  copper  chloride,  a  wetting  down  or  sprin- 
kling causes  an  additional  chlorination  of  3  to  6  per  cent.  Thus  at  the 
Lexington  mill,  Butte,  Mont.,  the  ore,  after  roasting  in  a  Stetefeldt  furnace, 
was  chloridized  to  65  per  cent,  after  two  hours  in  the  heap  to  75  or  80 
per  cent,  and  at  the  end  of  thirty-six  hours  to  92  per  cent  of  the  silver 
content. 

The  loss  in  silver  by  volatilization,  when  the  ore  has  been  properly  and 
carefully  roasted,  should  not  exceed  8  per  cent  except  in  presence  of 
volatile  elements  like  arsenic,  antimony,  selenium,  or  tellurium.  If,  how- 
ever, the  roasting  is  completed  at  a  high  temperature  the  loss  may  rise  to 
18  per  cent. 

Remarks  on  the  Chloridizing  Roast. — The  most  difficult,  and  at  the 
same  time  the  most  important  process  for  the  treatment  of  base  silver  ores 
by  wet  methods,  is  undoubtedly  chloridizing  roasting.  It  is  always  the 
safest  plan  for  the  operator  to  roast  as  thoroughly  as  possible.  If  the  ore 
is  well  chloridized,  sodium  hyposulphite  or  cyanide  extracts  all  the  silver 
chloride.  A  high  chloridization  does  not  necessarily  involve  a  high  loss  by 
volatilization.  It  is  well  suited  to  refractory  manganese  silver  ores,  as 


DRY  SILVER  MILLING  249 

most  of  the  silver  is  converted  into  a  readily  soluble  silver  chloride.  It 
has  never  been  favored  for  gold  ores  on  account  of  high  volatilization 
losses;  in  fact,  this  is  also  the  weakest  point  with  silver,  since  silver  chlorid 
is  quite  volatile. 

Chloridizing  by  Blast-roasting. — The  Dwight-Lloyd  machine,  Fig.  83, 
bids  fair  to  be  successfully  used  for  a  chloridizing  roast.  It  is  claimed  that 
the  volatilization  of  the  silver  is  entirely  under  control,  and  moreover  the 
cost  of  roasting  is  low. 

(3)     DRY  SILVER-MILLING  (REESE  RIVER  PROCESS) 

This  process  for  the  treatment  of  rebellious  silver  ores,  in  which  the 
metal  is  so  locked  up  as  to  require  roasting  before  it  can  be  amalgamated, 
was  developed  at  Reese  River,  near  the  Comstock  Lode  at  Virginia  City, 
Nev.  The  ore  contains  silver  sulphide,  particularly  the  antimonial  sul- 
phides, and  the  sulphide  of  the  base  metals  such  as  copper,  iron,  zinc,  and 
lead.  Galena,  however,  if  present  exceeding  5  to  10  per  cent,  renders  the 
ore  unsuitable  for  chloridization. 

The  treatment  in  brief  consists  in  dry-crushing  and  roasting  the  ore 
then  amalgamating  in  pans  to  recover  the  silver  and  gold.  The  dry- 
crushing  is  done  either  with  rolls  or  stamps.  Crushing  with  rolls  is  de- 
scribed in  the  chapter  on  Crushing.  If  dry-stamping  is  employed  the 
work  is  done  in  the  dry-crushing  silver  mill. 

(4)     THE  PATIO  PROCESS 

The  patio,  or  Mexican  amalgamation  process,  was  introduced  by 
Medina  into  Mexico  as  early  as  1557,  and  has  been  practiced  in  that 
country  down  to  the  present  time.  The  ores  best  suited  to  it  are  silicious 
ones  carrying  finely  disseminated  native  silver,  silver  sulphide,  and  chloride. 
A  limited  amount  of  pyrite,  galena,  cerussite,  or  the  copper  minerals  may 
be  present  without  serious  interference  with  the  process,  but  much  blende 
causes  low  extraction.  Where  any  gold  occurs  this  is  not  recovered. 

In  outline,  the  process  consists  in  finely  crushing  the  wet  ore,  and 
treating  the  mud  or  fine  product  in  a  flat  pile  in  a  large  paved  yard  or  patio, 
salt  and  bluestone  being  added  upon  the  pile  and  well  trodden  in  by  mules. 
Mercury  is  next  sprinkled  on  and  mixed  in  the  same  way.  The  above 
operations  require  two  to  four  weeks.  The  fine  product  is  then  washed 
in  tanks  to  separate  the  heavy  mercury  and  amalgam  from  the  light  tailing, 
and  the  amalgam  is  recovered  and  treated  as  in  silver  milling. 


CHAPTER  XXII 
SILVER  MILLING  BY  HYDROMETALLURGICAL  PROCESSES 

PRINCIPLES  OF  THE  HYDROMETALLURGY  OF  SILVER 

A  wet-process  for  the  recovery  from  the  ore  consists  in  dissolving  the 
metal  by  means  of  a  solvent  and  precipitating  from  the  solution  in  a  con- 
venient form.  The  silver  compounds  which  can  be  obtained  readily  in 
solution  are  the  sulphate  and  the  chloride.  In  cyanide  solution  argentite 
is  readily  soluble,  while  ruby  silver,  freislebenite,  and  stephanite,  are 
sparingly  so,  though  readily  soluble  in  mercurous  potassic  cyanide.  Silver 
sulphate  is  soluble  in  hot  water,  while  silver  potassic  chloride  is  .dissolved 
by  brine  solution  or  by  sodium  hyposulphite  (thiosulphate) .  From  the 
aqueous  solution  of  the  sulphate  silver  is  precipitated  by  metallic  copper; 
from  the  brine  solution  of  its  chloride  by  copper,  or  when  in  dilute  solution 
by  zinc,  iodide;  from  the  hyposulphite  solution  by  sodium  sulphide;  and 
from  the  cyanide  solution  by  metallic  zinc. 

The  Augustin  and  the  Ziervogel  processes,  introduced  in  1840  to  1850, 
were  used  in  a  limited  way  almost  exclusively  for  the  treatment  of  matte. 
However,  the  recent  application  of  the  Augustin  process  in  connection  with 
blast  roasting  for  the  treatment  of  low-grade  complex  silver  ores  containing 
lead  and  copper  is  worthy  of  note. 

THE  AUGUSTIN  PROCESS 

This  has  been  used  for  the  extraction  of  silver  from  ore  and  from 
copper-bearing  matte,  obtained  as  a  product  of  smelting.  At  Kosaka, 
Japan,  ore  consisting  of  one-half  heavy  spar  and  containing  10.5  oz.  silver 
per  ton  is  thus  treated.  The  ore  is  crushed  and  roasted  with  salt  in  a 
reverberatory  furnace,  and,  after  drawing  from  the  furnace  and  moistening 
on  the  cooling-floor,  contains  80  per  cent  of  the  silver  in  the  form  of  chloride. 
It  is  leached  with  a  hot  18  per  cent  salt  solution  in  regular  leaching-vats. 
The  leaching  is  continued  until  a  polished  plate  of  copper  shows  no  precip- 
itate of  silver  when  held  in  the  flowing  filtrate.  It  requires  0.66  ton  of 
brine  to  leach  a  ton  of  the  ore.  The  sand  is  washed  with  hot  water,  and 
the  tailing  rejected. 

250 


THE  ZIERVOGEL  PROCESS 


251 


THE  ZIERVOGEL  PROCESS 

This  process,  practiced  at  Mansfeldt,  Germany,  and  at  the  Boston  & 
Colorado  smelting  works  at  Argo,  Colo.,  is  adapted  to  the  treatment  of 
rich  copper  matte  containing  little  or  no  arsenic,  antimony,  or  bismuth, 
any  of  which  would  form  insoluble  compounds  with  silver.  The  method 
may  be  divided  into  three  parts:  the  roasting  for  silver  sulphate,  the  leach- 
ing, and  the  precipitation  of  the  silver. 

The  Process. — Referring  to  the  flow-sheet  of  the  process  (see  Fig.  140) 


|< Copper  Ores 

U Silver  &  Gold  Ores 


Silicious  Copper  Ore 
Sulphide      •• 


Water 


Plate* 


Final  Solution 
to  Waste 


FIG.  140. — Flow-sheet  of  Ziervogel  Process. 

we  have  in  furnaces  A,  the  operation  of  producing  the  matte  or  regulusfrom 
gold-  and  silver-bearing  copper  ores.  The  details  of  the  process  are 
described  in  the  chapter  on  the  Metallurgy  of  Copper,  under  the  head  of 
"  Reverberatory  Matte  Smelting."  The  composition  of  the  matte  is  Cu, 
47.3  per  cent;  Pb,  8.1;  Zn,  2.7;  Fe,  17.7;  S,  21.6  with  400  oz.  silver  and 
15  oz.  gold  per  ton. 

Preparation  of  the  Matte. — The  matte  is  crushed  and  passed  through 
rolls  at  B  to  reduce  it  to  6^-mesh  size,  and  sent  to  a  reverberatory  furnace  C, 
where  it  receives  preliminary  roasting.  The  roasting  reduces  the  sulphur 
to  6.3  per  cent,  and  converts  the  iron  and  copper  sulphides  to  the  corre- 
sponding oxides  and  sulphates,  as  described  in  the  chapter  on  the  chemistry 


252       SILVER  MILLING  BY  HYDROMETALLURGICAL  PROCESSES 

of  Oxidizing  Roasting.  This  partly  roasted  product  then  goes  to  a  Chilian 
mill  D  (see  also  Fig.  46),  where  it  is  finely  ground  to  60-mesh. 

Sulphatizing  Roasting. — The  partly  roasted  matte  is  next  treated  by 
hand  in  charges  of  1600  Ib.  by  a  sulphatizing  roast  in  small  single-hearth 
reverberatory  roasters  at  E.  In  the  process  the  iron  and  copper  remaining 
in  the  form  of  sulphides  are  converted  into  sulphates  which  react  on  the 
silver  sulphide  at  a  slightly  higher  temperature,  as  follows: 

(10)  Ag2S+3O+CuSO4  =  Ag2SO4+CuO+SO2. 

It  has  been  found  that  the  addition  of  2  per  cent  sodium  sulphate  (salt 
cake)  facilitates  the  change.  The  roasting  takes  place  in  four  stages  as 
shown  below. 

During  the  first  stage,  of  1J  hours,  the  draft  is  checked,  the  side  doors 
kept  open,  and  the  charge  held  at  a  low  temperature.  The  charge  becomes 
evenly  heated  throughout,  and  glows  from  the  oxidation  of  Cu2S  to  Cu2O. 

During  the  second  stage,  of  1 J  hours,  the  heat  is  increased  and  the  charge 
constantly  rabbled.  Iron  sulphate  is  decomposed  with  the  consequent 
formation  of  copper  sulphate.  The  charge  swells  and  becomes  spongy  by 
the  formation  of  this  salt. 

In  the  third  stage,  the  temperature  is  increased  for  an  hour  until  tests 
show  that  the  silver  is  "  out,"  that  is,  in  the  form  of  sulphate.  The  follow- 
ing reaction  occurs: 

(11)  CuSO4+Ag2O  =  Ag2S04+CuO. 

During  the  fourth  stage  the  temperature  is  kept  constant.  The 
charge  is  gathered  and  pressed  down  with  a  heavy,  long-handled  iron  paddle 
to  break  the  lumps,  and  then  vigorously  stirred  to  oxidize  the  remaining 
Cu2O  to  CuO,  and  decompose  copper  sulphate.  The  temperature  is  not 
further  increased,  since  it  would  decompose  silver  sulphate,  forming  silver 
oxide,  rendering  the  silver  again  insoluble. 

The  progress  of  the  roast  is  tested  by  dropping  small  samples  from  time 
to  time  into  hot  water.  Soluble  sulphates  dissolve  in  the  hot  water;  and 
in  the  tests  made  early  the  solution  becomes  deep  blue.  Later,  as  the  silver 
sulphate  begins  to  form,  it  is  immediately  reduced  to  silver  spangles  uy  the 
cuprous  oxide  present.  As  the  roasting  advances  during  this  stage,  the 
copper  sulphate  decomposes,  and  the  solution  becomes  less  blue  in  the  test 
and  the  silver  spangles  increase  and  afterward  diminish.  During  the 
fourth  stage  the  Cu2O  is  changed  to  CuO  and  th'e  spangles  no  longer  show. 
A  light-blue  color  of  the  solution  remains,  due  to  the  presence  of  a  little 
copper  sulphate,  which  indicates  that  the  silver  sulphate  is  not  itself 
becoming  decomposed.  A  sample  thus  roasted  showed  by  analysis 


THE  2IERVOGEL  PROCESS  253 

2.5  per  cent  FeSO4  and  ZnSO4;  0.6  per  cent  CuSO4,  and  1.73  per  cent 
Ag2SO4  (348  oz.  silver  per  ton),  so  that  there  was  left  in  the  matte  (there 
being  no  loss  of  weight  in  roasting  matte)  52  oz.  per  ton  or  13  per  cent  of 
the  silver  in  insoluble  form. 

Leaching. — The  roasted  matte  is  charged  into  tanks  F  and  leached 
with  hot  water  to  dissolve  the  sulphate  above  described.  The  filtrate  goes 
to  a  series  of  boxes  H,  containing  copper  plates  upon  which  the  silver  pre- 
cipitates in  the  form  of  white  shining  crystals.  The  silver-free  solution, 
containing  in  addition  to  the  original  copper  sulphate  that  which  it  has 
taken  from  the  copper  plates,  goes  to  tanks  7,  where  the  copper  is  precip- 
itated upon  scrap-iron  to  recover  the  copper.  The  final  solution  is  rejected. 

The  cement-silver  from  the  precipitating  boxes  is  transferred  to  a  tank, 
and  dilute  sulphuric  acid  is  added.  It  is  boiled  by  forcing  in  a  mixture  of 
air  and  steam  from  an  injector.  The  treatment  oxidizes  and  dissolves  the 
traces  of  copper  still  retained  by  the  silver  crystals,  and  keeps  them  in 
agitation  at  the  boiling  temperature  of  the  acid  mixture.  The  copper  sul- 
phate solution  is  now  run  off  and  the  residue  repeatedly  washed  by  decan- 
tation  with  hot  water  to  free  it  entirely  from  copper.  It  is  transferred  to  a 
long  pan  over  a  coal  fire  for  drying  and  is  then  melted  down  in  crucibles 
in  a  wind-furnace  and  obtained  in  ingots  999  to  999.5  fine. 

Residue  from  the  Leaching  Tanks. — The  extracted  residue  remaining 
in  the  tanks  F,  still  retaining  52  oz.  silver  per  ton  as  above  stated,  freed 
from  sulphates,  and  composed  mainly  of  iron  and  copper  oxides,  is  sent  to  a 
reverberatory  furnace  K,  to  form  copper  matte.  The  slag  produced  in  the 
treatment  goes  back  to  the  ore-smelting  furnace  A ,  while  the  matte,  tapped 
into  sand  molds,  is  sent  to  the  reverberatory  furnace  L,  to  be  treated  by 
the  English  process  of  making  "  best-selected  copper."  Here  the  matte, 
in  large  lumps,  is  piled  up  in  the  furnace  near  the  bridge  and  exposed  to 
a  flame  made  oxidizing  by  an  excess  of  air  admitted  through  the  fire  and 
through  openings  in  the  bridge  and  roof  of  the  furnace.  The  effect  is 
to  "  roast  "  the  matte  as  the  lumps  slowly  melt  and  the  drops  of  liquefying 
matte  come  in  contact  with  the  air.  Finally  the  whole  charge  becomes 
melted,  and  the  copper  oxide  which  has  been  formed,  acts  on  the  unoxidized 
copper  sulphide  of  the  matte  as  follows: 

(12)  2Cu2O+Cu2S  =  6Cu+SO2. 

The  aim  is  to  extend  the  roasting  only  so  far  as  to  obtain  in  the  form  of 
metallic  copper  one-fifteenth  of  the  total  matte.  When  the  charge  is 
tapped  from  the  furnace  into  molds,  made  in  the  sand,  the  copper  is  found 
in  the  form  of  plates  or  bottoms  in  the  first  of  the  molds  beneath  the  lighter 
matte.  The  bottoms  absorb  the  impurities  such  as  arsenic,  antimony, 
lead,  and  bismuth,  practically  all  the  gold  (100  to  200  oz.  per  ton)  and 


254        SILVER  MILLING  BY  HYDROMETALLURGICAL  PROCESSES 

some  of  the  silver.  On  the  other  hand,  the  supernatant  matte  has  risen 
to  the  grade  of  white-metal  of  75  per  cent  copper,  and  carries  90  to  100  oz. 
silver  but  not  more  than  0.2  oz.  gold  per  ton. 

To  prepare  it  for  the  extraction  of  the  silver,  the  matte  is  again  given  a 
sulphatizing  roast,  but  in  a  different  furnace  from  the  one  used  for  the  first 
matte.  The  residue  after  this  second  treatment,  principally  a  copper  oxide 
containing  10  oz.  silver  per  ton,  is  sold  to  the  oil  refiners.  The  bottoms, 
formerly  treated  at  the  Argo  works  by  a  secret  process  for  the  extraction 
of  the  precious  metals,  is  now  elect rolytically  refined. 

THE  HYPOSULPHITE  LIXIVIATION  OF  SILVER  ORE  (VON  PATERA 

PROCESS) 

Hyposulphite  lixiviation  can  be  practiced  upon  ore  containing  simple  or 
compound  sulphides  of  silver  that  have  undergone  a  preliminary  chloridizing 
roasting.  The  silver  sulphides,  in  the  roasting,  become  converted  into 
silver  chloride.  The  process  also  applies  to  silver  ore  already  containing 
the  silver  as  chloride.  Free-milling  ore,  such  as  oxidized  ore  containing 
the  silver  in  the  native  state,  or  as  chloride,  or  to  some  extent  as  argentite, 
are  preferably  treated  by  milling  and  amalgamation.  Native  silver  and 
silver  sulphide  in  a  favorable  form  can  be  recovered  by  milling  and  amal- 
gamation, whereas  by  hyposulphite  extraction,  they  would  remain  insoluble. 
The  process  is  not  suited  to  the  treatment  of  gold  ore.  The  extraction  of 
gold  is  low;  usually  less  than  50  to  70  per  cent.  One  of  the  most  useful 
applications  is  to  the  treatment  of  argentiferous  blende  that  has  been 
hand-picked  or  concentrated  from  galena,  and  may  still  contain  lead  up  to 
8  per  cent. 

The  hyposulphite  process  is  based  upon  the  fact  that  silver  chloride 
readily  dissolves  in  dilute  solutions  of  sodium  hyposulphite.  The  chlorid- 
izing roasting  is  unquestionably  the  most  important  part  of  the  process, 
and  the  chief  attention  and  study  is  to.be  given  it. 

It  consists  in  crushing  the  ore,  roasting  it,  and  treating  the  roasted  ore 
in  filter-bottom  vats,  first  with  water  to  remove  the  soluble  chlorides  and 
sulphates  of  the  heavy  metals  (base  metal  leaching),  then  leaching  with  a 
dilute  solution  of  sodium  hyposulphite  to  dissolve  the  silver  chloride. 
Silver  sulphide  is  precipitated  from  the  filtrate  with  sodium  sulphide,  cfyied, 
and  roasted  to  remove  the  sulphur,  and  the  residue  is  sent  to  the  smelting 
works,  or  treated  in  an  English  cupelling  furnace. 

THE  RUSSELL  PROCESS 

This  is  a  modification  of  the  Patera  process,  principally  by  the  use  of 
another  solution,  in  addition  to  the  hypo-solution,  for  the  extraction  of  the 
silver.  By  mixing  in  solution  two  parts  of  the  hypo-salt  with  one  of  copper 


THE  RUSSELL  PROCESS  255 

sulphate  we  obtain  a  double  salt  ^28263  •  Cu2Oa,  called  the  extra-solution. 
It  has  a  solvent  power  nine  times  as  great  as  that  of  the  ordinary  hypo- 
solution  for  native  silver,  silver  sulphides,  silver  arsenides,  and  silver  anti- 
monides.  In  the  case  of  an  imperfectly  roasted  ore  the  use  of  the  extra- 
solution  insures  the  extraction  of  more  silver  from  the  compounds  men- 
tioned than  could  be  obtained  by  the  use  of  ordinary  hypo-solution. 


CHAPTER  XXIII 
CYANIDATION  OF  SILVER  ORES 

PRINCIPLES  OF  CYANIDATION 

Silver  ores  carrying  gold,  and  in  which  the  silver  occurs  as  chloride,  or 
argentite,  or  stephanite,  have  been  successfully  cyanided.  Manganese 
silver  ores  contain  the  silver  intimately  associated  with  the  manganese 
and  will  give  no  extraction  when  treated  raw,  unless  there  be  a  preliminary 
acid  treatment. 

Of  the  silver  minerals,  native  silver,  in  particles  so  large  as  to  be  visible, 
is  insoluble  in  potassium  cyanide  in  any  reasonable  time.  Silver  chloride, 
bromide,  arid  argentite,  are  readily  soluble.  Ruby  silver,  stephanite,  and 
freieslebenite  are  sparingly  soluble  in  potassium  cyanide,  but  readily  sol- 
uble in  mercurous  potassium  cyanide  solution. 

A  chloridizing  roast  is  always  beneficial  to  ores  containing  silver  sul- 
phides as  it  increases  the  percentage  of  extraction.  However,  it  is  not 
essential  and  is  seldom  employed  in  the  cyanide  treatment  of  silver  ores. 

In  handling  silver  ores  the  reactions  are  more  complex  than  in  treating 
ores  of  gold,  owing  to  the  greater  chemical  activity  of  the  silver  compounds, 
and  to  the  fact  that  owing  to  the  larger  quantity  of  contained  metal  the 
cyanide  solutions  are  necessarily  stronger. 

Important  matters  in  cyaniding  silver  ore  are  the  following : 

(a)  A  long  time,  ten  to  twenty-five  days  in  the  case  of  sand  treatment, 
is  needed  for  leaching.  For  slime  treatment  from  forty-eight  to  ninety-six 
hours  would  suffice  for  a  complete  cycle,  in  which  time  a  higher  percentage 
of  extraction  would  be  obtained  than  by  a  fourteen-day  treatment  of  the 
corresponding  sand.  The  silver  compounds  are  not  so  easily  soluble  as 
gold,  and  a  larger  amount  must  be  dissolved. 

(6)  Thorough  oxygenation  is  necessary,  not  only  because  of  the  large 
amount  of  silver  present,  but  because  silver  compounds  need  at  least 
initial  oxidation  to  become  properly  soluble  in  cyanide  solution.  Hence 
an  advantage  is  secured  by  the  double  treatment  of  sand.  Also  during 
leaching,  if  the  solution  be  allowed  to  sink  several  inches  below  the  top  of 
the  charge,  before  another  wash-solution  is  run  on,  air  is  drawn  in  and 
penetrates  the  ore,  and  the  solution  following  forces  the  air  downward 
through  the  ore.  In  the  treatment  of  the  slime  the  pulp  may  receive  thor- 
ough aeration  by  agitating  with  air. 

256 


PRECIPITATION  OF  THE  SILVER  257 

(c)  Stronger  solution  is  used  than  for  the  treatment  of  gold  ore.     Thus 
the  first  or  strong  solution  may  be  0.7  per  cent,  the  weak  one  0.25  per  cent, 
while  for  gold  ore,  a  0.5  per  cent  solution  would  be  called  strong  and  0.05 
per  cent  weak. 

(d)  The  consumption  of  potassium  cyanide  is  higher  than  in  the 
treatment  of  gold  ores.     It  varies  from  1.5  to  4  Ib.  per  ton  as  compared 
with  0.4  to  0.8  Ib.  consumed  in  the  treatment  of  gold  ores. 

(e)  The  precipitation  of  silver  from  cyanide  solution  by  zinc  shaving 
presents  no  difficulties  and  is  practically  complete.     Despite  the  fact  that 
a  relatively  great  amount  of  silver  has  to  be  precipitated,  as  compared  with 
gold,  no  more  zinc  is  consumed. 

THE  PRECIPITATION  OF  SILVER  FROM  CYANIDE  SOLUTION 

Silver  in  cyanide  solution  may  be  recovered  by  precipitating  upon  zinc 
shaving  or  by  zinc  or  aluminum  dust  according  to  the  reactions  (6)  and  (8), 
page  147.  In  presence  of  copper  the  precautions  described  under  head  of 
"  Precipitation  of  Gold  from  Cyanide  Solutions  "  will  equally  apply  to 
silver  precipitation.  Silver  ores  of  course  make  a  much  larger  bulk  of 
precipitate  than  gold.  On  this  account  the  Merrill  precipitation  process  is 
preferred  to  the  use  of  the  zinc  boxes  illustrated  on  page  258,  though 
for  a  small  plant  they  might  be  used. 

Fig.  141  shows  in  outline  the  course  of  procedure  for  the  clean-up,  the 
pressing,  drying,  and  melting  of  the  precipitate  from  clarified  silver  solu- 
tions. The  novel  feature  is  the  washing  of  the  fine  precipitate  from  the 
zinc  box  with  pregnant  solution  into  the  filter  press,  where  it  is  collected. 
The  pumping  of  pregnant  solution  through  the  filter  press  is  continued  at 
intervals,  until  the  tail  solution  rises  to  approximately  the  value  of  the 
pregnant  solution  precipitated.  During  the  earlier  stages  of  pumping, 
the  tail  solution  from  the  filter  press  is  of  low  enough  grade  to  divert  to 
the  milling  solution,  but,  during  the  later  stages,  it  is  advisable  to  return 
it  to  the  head  of  the  zinc  boxes.  With  solutions  in  which  silver  occurs 
in  the  ratio  of  50  parts  to  1  part  of  gold,  precipitate  is  at  times  obtained 
which  contains  approximately  90  per  cent  of  silver  and  gold.  There  is 
no  difficulty  in  obtaining  precipitate,  regularly,  well  over  80  per  cent  of 
silver  and  gold.  When  of  this  grade  it  is  very  easily  melted,  with  a 
minimum  consumption  of  fluxes  and  fuel,  and,  of  course,  produces  a  high- 
grade  bullion.  At  a  certain  Mexican  plant,  where  this  plan  was  adopted, 
the  grade  of  the  precipitate  was  very  materially  raised  and  the  fineness  of 
the  bullion  was  increased  from  800  to  900.  This  resulted  in  a  monetary 
saving  of  several  hundred  dollars  per  month.  A  similar  plan  appears  to 
be  feasible  with  zinc  dust  precipitation.  It  is  still  a  question  as  to  how 
far  such  a  plan  could  be  carried  with  solutions  in  which  gold  predominates. 


258 


CYANIDATION  OF  SILVER  ORES 


With  the  Merrill  process  double  precipitation  is  generally  practiced,  as 
shown  in  Fig.  117;   the  apparatus  is  in  duplicate,  constituting  two  circuits. 


In  the  one  circuit  the  heavier  feed  of  zinc-dust  is  maintained  to  give  a 
barren  solution,  while  in  the  other  there  is  just  sufficient  to  precipitate  the 
bulk  of  the  silver.  The  amount  of  the  barren  solution  to  be  made  is 


THE  SANTA  GERTRUDIS  REFINERY  259 

determined  by  the  needs  of  the  final  washes,  while  the  other,  needing  less 
zinc-dust  and  carrying  a  few  cents  worth  of  silver,  is  used  in  the  agitation 
tanks.  Thus  in  a  Mexican  silver-mill  of  500  tons  capacity,  about  half  the 
solution  was  completely  precipitated  to  retain  but  1  to  2  cents  worth  of 
silver  per  ton,  using  1.1  oz.  of  zinc-dust  to  one  of  silver.  On  the  other 
hand  the  partially  or  almost  precipitated  solution  needed  but  0.74  oz.  of 
dust  to  one  of  silver  and  this  retained  but  10  cents  of  solution,  well  suited 
for  re-use. 

THE   SANTA  GERTRUDIS  PRECIPITATING  AND   REFINING  PLANT 

Fig.  142  is  a  plan  of  this  installation,  showing  the  method  used  for 
double  precipitation  by  the  Merrill  process. 

There  are  four  pregnant  solution  storage  tanks,  so  that  an  exactly 
measured  quantity  of  solution  can  be  precipitated  by  a  known  weight 
of  zinc-dust.  Owing  to  the  large  quantity  of  zinc-dust  needed  this 
is  slowly  fed  to  the  zinc-dust  mixing  cone  by  an  endless  belt  through 
a  worm-gear  drive,  this  belt  taking  its  feed  from  a  supply-hopper 
in  regulated  quantity.  Referring  now  to  the  partial  precipitation  circuit, 
the  two  tank-discharges  unite  in  one  just  before  the  zinc-feeder.  The 
emulsion  from  the  mixing  cone  drops  by  a  vertical  pipe  into  the  tank 
discharge  which  constitutes  the  12-in.  suction  of  pumps  Nos.  1  and  2  of  the 
circuit.  From  the  pumps  the  solution  now  impregnated  with  the  zinc- 
dust,  passes  to  the  Merrill  presses  Nos.  1,  2,  3,  and  4.  When  the  desired 
tank,  such  as  tank  No.  9,  is  emptied  and  the  solution,  now  nearly  freed 
from  its  silver,  is  filtered  and  washed,  the  filtrate  goes  by  a  10-in.  pipe  line 
to  mill-solution  sump-tank  No.  15.  The  emptied  presses  are  blown  with 
air  to  drop  off  the  precipitate  on  the  filter  leaves  and  the  press  opened  for 
the  removal  of  the  precipitate.  For  the  production  of  barren  solution  a 
similar  procedure  is  followed,  taking  the  pregnant  solution  from  tanks  Nos. 
11  and  12  and  delivering  through  presses  5,  6,  and  7  to  the  barren  solution 
sump-tank  No.  14. 

Fig.  115  is  a  view  of  the  style  of  press  used.  The  filter  press  frames  are 
triangular,  the  pulp  entering  at  the  bottom  of  the  frame.  The  solids  grad- 
ually accumulate  on  the  filter  cloths  of  the  frames,  providing  a  uniform 
filtering  layer  of  a  fine-grained  precipitant  through  which  the  solution 
must  pass  to  ensure  contact  between  the  zinc  and  the  gold  particles  for 
instant  precipitation.  The  barren  or  nearly  barren  solution,  the  product 
of  the  press,  then  passes  by  way  of  the  discharge  launders  to  the  completed 
precipitate  or  to  the  partially  precipitated  solution  tank. 

For  a  clean-up  the  press-cakes  are  dried  by  forcing  air  through  the 
precipitate  on  the  filters,  the  press  is  opened  and  the  caked  material  dumped 
into  tram-cars.  The  product  is  sampled,  weighed  and  determined  for 
contained  moisture,  thus  giving  its  exact  dry  weight. 


260 


CYANIDATION  OF  SILVER  ORES 


We  give  in  Figs.  1 16  and  1 17  views  of  a  press-frame  and  of  a  press-plate. 
There  are  say  forty  such  frames  in  a  press.  The  plate  is  covered  on  both 
of  its  faces  by  a  filter  cloth.  When  the  frames  and  plates  are  placed  tightly 


together  in  the  press  the  common  circular  passage  along  the  top  admits 
the  solution  to  each  frame.  It  is  led  to  the  bottom  of  the  frame  by  the 
drop  pipe  there  shown.  Passing  through  the  filter  cloth  it  escapes  by  a 
cock  at  the  left-hand  upper  corner  of  the  plate. 


REFINING  SILVER  PRECIPITATE  261 


PRECIPITATION  BY  ALUMINUM  DUST 

This  is  not  new  as  far  as  the  actual  knowledge  of  its  use  is  concerned; 
but  it  was  only  recently  that  its  use  in  a  practical  way  on  a  large  scale 
was  studied.  This  is  done  at  Deloro  plant  and  the  Nipissing  mines, 
Ontario,  where  rich  silver  ore  containing  arsenic  and  cobalt  is  treated. 
They  are  perhaps  exceptional  cases,  rendering  its  use  necessary. 

Precipitation  at  Nipissing. — It  was  found  here  that,  after  solutions 
had  been  precipitated  by  zinc  they  rapidly  lost  their  dissolving  efficiency, 
so  experiments  led  to  the  use  of  aluminum  dust.  The  metal  must  be  well 
agitated  with  solutions  before  precipitation.  The  fact  that  aluminum  does 
not  replace  the  precious  metals  in  the  cyanogen  compound  renders 
necessary  the  presence  of  a  caustic  alkali.  The  reaction  in  precipitation 
is  probably 

(13)      6NaAgCN2+6NaOH+Al  =  6Ag+12NaCN+2Al(OH)3. 

From  this  equation  it  will  be  seen  that  there  is  a  regeneration  of  cyanide. 

At  this  plant  the  pregnant  solution  is  pumped  to  a  sand-filter.  It  is 
then  precipitated  by  the  Merrill  process  as  above  described.  Some  550  to 
600  tons  of  solution  are  handled  daily,  the  head-assay  running  about  8.25 
oz.  the  tail  assay  or  barren  solution  0.10  oz.  of  silver  per  ton.  Precipita- 
tion averages  98  per  cent,  using  0.02  Ib.  of  aluminum  per  ounce  of  silver 
at  a  cost  of  18.5  cents  per  ton  of  ore  treated. 

DRYING  AND  REFINING  SILVER  PRECIPITATE 

The  early  practice  of  selling  the  precipitate  to  smelters  is  still  adhered  to 
by  a  few  companies,  but  the  majority  convert  it  into  bullion  before  market- 
ing. Cyanide  bullion  varies  greatly  in  fineness,  depending  upon  the  char- 
acter of  the  ore  treated  as  well  as  the  equipment  provided  for  refining  and 
the  degree  of  skill  exercised  in  its  use.  It  is,  obviously,  not  good  practice 
to  carry  local  refining  to  the  point  where  the  cost  is  greater  than  the  advan- 
tages to  be  realized  from  marketing  higher-grade  bullion.  The  only  case 
that  I  know  of  where  fine  bullion  is  produced  which  requires  no  further 
refining  is  that  of  the  Nipissing  Mining  Company,  Cobalt,  Ont.,  Canada. 
Here,  unusual  conditions  make  the  production  of  fine  bullion  a  com- 
paratively easy  matter. 

Acid  Treatment. — Preliminary  acid  treatment  of  the  precipitate  has 
been  found  unnecessary  in  most  cases  where  silver  predominates,  but  is 
still  adhered  to  in  the  majority  of  cases  for  gold  precipitate. 

The  Tavener  Method. — The  Tavener  method  of  refining,  involving  the 
melting  of  the  precipitate  in  a  small  reverberatory  furnace  with  various 
fluxes  and  lead,  followed  by  cupellation,  is  still  in  general  use  in  South 
Africa,  but  has  not  gained  ground  in  this  country. 


262 


CYANIDATION  OF  SILVER  ORES 


Cupelling. — The  Homestake  method  of  refining,  involving  acid  treat- 
ment and  briquetting  with  lead  flux,  followed  by  melting  and  cupellation 
in  the  ordinary  English  cupel  furnace,  as  well  as  the  treating  of  all  the  by- 
products in  a  small  blast  furnace,  is  used  by  a  few  American  plants. 

Blast-furnace  Treatment. — The  melting  of  the  precipitate,  briquetted 
with  the  proper  fluxes,  in  a  small  blast-furnace,  followed  by  cupellation  of 
the  lead,  has  been  found  of  advantage  when  dealing  with  a  large  volume  of 
low-grade  precipitate.  Losses  from  dusting,  which  might  appear  to 
be  the  chief  objection  to  this  method  of  melting,  are  claimed  to  be 
insignificant  when  a  proper  flux  system  is  provided.  This  method 
of  melting  is  more  economical  than  either  the  Tavener  or  Homestake 
practice. 

The  Electric  Furnace. — The  electric  furnace  has  also  been  used  in  a 
few  cases. 

The  Tilting  Furnace. — The  most  simple  and  satisfactory  method  of 
converting  the  high-grade  precipitate  which  is  obtained,  by  proper  manip- 
ulation, from  ores  in  which  silver  predominates  is  to  melt  the  precipi- 
tate directly,  with  the  minimum  proportion  of  flux,  in  the  tilting  type  of 
furnace.  A  few  mills  use  a  double-chamber  tilting  furnace,  in  which  the 
flame  comes  in  direct  contact  with  the  charge  being  melted.  This  furnace 
is  perhaps  more  economical  of  fuel,  but,  unless  the  precipitate  is  briquetted 
there  is  greater  risk  of  loss  through  dusting  than  when  the  precipitate 
is  melted  in  a  closed  crucible.  In  the  case  of  silver  precipitate,  with 
proper  manipulation  in  the  crucible  furnace,  it  is  a  serious  question 
whether  it  pays  to  briquette,  the  losses  being  less  under  these  conditions 
than  the  cost  of  briquetting. 

Let  us  take  the  case  of  a  low-grade  precipitate  containing  25  per 
cent  gold  and  silver;  39  per  cent  lead,  20  per  cent  zinc,  J  per  cent  copper; 
\  per  cent  sulphur,  3  per  cent  lime  and  1.4  per  cent  silica.  In  the  fol- 
lowing table  we  will  have: 


SLAG  BALANCE. 

OXYGEN. 

Substance. 

Weight,  Pounds. 

Combined  as. 

Acid,  Pounds. 

Base,  Pounds. 

Zn 

20.60 

37.50 
0.27 
2.30 
19.00 
27.00 
22.50 

ZnO 
PbO 
Cu2O 
CaO 
Si02 
B2O3 
Na2O 

5.10 
2.90 
0.03 
0.66 

5.80 

Pb 

Cu...  

CaO  
SiO2                  .'..'. 

10.2 

18.5 

*B2O< 

*Na2O  

28.7 

14.49 

*  Computed  from  weights  of  borax  and  niter  used. 


REFINING  SILVER  PRECIPITATE 


263 


Here  the  ratio  of  acid  to  basic  oxygen  is  as  2  to  1.  There  was  yielded  134 
Ib.  of  slag  and  5J  Ib.  of  a  white-looking  layer  or  cover  of  sulphate  of  soda. 
Of  niter  there  was  used  35  Ib.,  of  which  6  Ib.  was  needed  for  the  cover,  leav- 
ing 29  Ib.  to  give  the  22.5  Ib.  of  soda  for  the  slag.  Of  borax  39  Ib.  was 
required  to  yield  27  Ib.  of  I^Os,  so  that  there  was  a  total  of  93  Ib.  of  fluxes 
for  100  Ib.  of  precipitates.  When  melted  the  resultant  bullion  was  972 
fine.  A  plumbago  crucible  is  preferred,  but  where  there  is  much  copper 
there  is  a  clay  crucible  used  because  a  plumbago  crucible  would  tend  to 
reduce  the  copper  which  would  then  enter  the  bullion,  reducing  its  fineness. 
Types  of  Furnaces  Used. — Various  types  of  furnaces  are  used  in  melting 
precipitate,  namely,  the  ordinary  wind-furnace  holding  from  a  60-  to  200- 
size  crucible  (see  Fig.  1);  the  Faber-du  Faur  tilting  furnace  as  used  at 


FIG.  146. — Monarch  Rockwell  Refining  Furnace. 

Kalgoorlie;  the  Steele-Harvey  oil-fired  tilting  furnace;  the  Monarch- 
Rockwell  as  used  at  the  Belmont,  Tonopah,  the  latter  as  shown  in  Fig.  146. 
Drying,  Melting  and  Refining  by  Fusion. — The  damp  precipitate  from 
the  ore  is  well  mixed  with  a  calculated  amount  of  fluxes,  the  moisture 
loaded  into  pans  and  placed  in  a  large  muffle  furnace  and  gradually 
heated  to  a  red  heat.  Moisture  is  driven  off  and  the  precipitate  in 
the  acid  settles  to  about  one-third  of  its  original  bulk.  The  sintered  mass, 
free  from  dust,  gives  an  excellent  product  for  crucible  melting.  The  fluxes 
used  are  generally  Chili  niter  (sodium  nitrate)  for  oxidizing  soda-ash  to 
take  up  the  sulphur,  borax,  and  sand  as  acid  fluxes  for  the  bases.  These 
are  added  in  such  proportions  as  to  produce  with  the  silica  and  the  bases 
in  the  precipitate  a  slag  of  two  of  acid  to  one  of  base  oxygen.  The  ratio 


264  CYANIDATION  OF  SILVER  ORES 

of  borax  glass  to  silica  should  be  as  2  to  1 ;  or  in  presence  of  much  lead  as 
1  to  1. 

Smelting  in  an  Oil-burning  Reverberatory  Furnace. — The  hearth  of  the 
furnace  is  11  ft.  by  4  ft.  9  in.  wide  and  will  take  a  charge  of  5  tons  of  pre- 
cipitate. Before  using,  the  furnace  is  seasoned  by  melting  in  700  Ib.  of 
old  slag  which  consolidates  the  bottom.  The  precipitate  is  properly 
fluxed  and  melted  down,  giving  about  2J  tons  of  bullion  of  883  fine  in  silver. 
The  furnace  is  operated  as  often  as  a  5-ton  charge  is  ready,  the  smelting 
taking  from  sixteen  to  twenty-eight  hours. 

CHEMISTRY  OF  THE  CYANIDE  PROCESS  FOR  SILVER  ORES 

When  a  solution  of  potassium  cyanide  is  brought  in  contact  with  finely 
divided  silver  or  silver  chloride  the  metal  is  dissolved  according  to  the 
following  equation : 

(14)  2Ag+4KCN+O+H2O  =  2AgK(CN)2+2KOH. 

The  reaction  is  similar  when  the  cheaper  sodium  sulphide  is  used. 

Oxygen  is  needed  to  complete  the  reaction,  so  that  the  ore-pulp  should 
be  properly  aerated.  This  may  take  many  hours.  When  the  occluded 
oxygen  is  used  up  the  reaction  ceases,  but  resumes  with  a  fresh  supply  of 
air.  As  described  for  gold  ferrous  and  ferric  sulphates  result  from  the 
decomposition  of  pyrites  and  tend  to  precipitate  silver.  To  overcome 
this  an  addition  of  quicklime  is  made  in  more  than  sufficient  quantity  to 
overcome  the  acidity,  the  excess  being  termed  "  protective  alkalinity." 
When  silver-bearing  sodium  cyanide  solution  is  brought  in  contact  with 
zinc  shavings  or  zinc  dust  we  have  the  reaction : 

(15)  NaAg(CN)2+2NaCN-fZn+H2O  =  NaZn(CN)4+Ag+H+NaOH. 

One  part  of  zinc  is  calculated  to  precipitate  1.7  parts  of  silver.  But 
cyanide  is  also  used  up  by  direct  combination  with  the  zinc. 

(16)  Zn+4KCN+2H2O  =  K2Zn(CN)4+2KOH+H2. 

In  both  these  reactions  there  is  an  escape  of  hydrogen  bubbles.          t 
When  aluminum  is  used  as  precipitant  we  have : 

(17)2NaAg(CN)2+4NaOHH-2Al  =  4NaCN+2Ag-fNa2Al204+4Hi, 

in  which  one  part  of  aluminum  precipitates  four  parts  of  silver,  in  this 
respect  being  so  much  more  efficient  than  zinc.1 

1  In  place  of  writing  sodium  cyanide  NaCN  it  is  often  written  NaCy,  the  Cy  being  a 
convenient  way  of  specifying  the  CN. 


THE  BELMONT  MILL  265 

TYPICAL  SILVER  MILLS 
The  following  are  descriptions  of  silver  mills: 

BELMONT  MILLING  CO.  MILL,  TONOPAH,  NEVADA 

The  ore  is  a  fine  granular  quartz  of  about  72  per  cent  silica  with  a  few 
per  cent  of  sulphides.  The  silver  occurs  as  disseminated  silver  sulphides, 
antimonides,  and  selinides,  a  little  native  silver,  and  a  little  gold.  Thus 
to  0.32  oz.  of  gold  there  would  be  32  oz.  silver,  or  in  weight  1  of  gold  to  100 
of  silver.  The  ore  is  treated  by  concentration  and  the  tailings  agitated  in 
cyanide  solution  and  filtered.  The  operations  may  be  divided  into  five; 
(1)  picking,  coarse  crushing^  and  delivery  to  the  mill  bins;  (2)  stamping, 
tube-milling,  and  concentrating  to  remove  the  heavy  part  to  be  sold  to  the 
smelter;  (3)  cyaniding  the  tailings  from  the  concentrating  tables  to  yield 
a  pregnant  solution;  (4)  precipitation  of  the  gold  from  the  solution;  (5) 
refining  the  precipitate. 

(1)  Picking,  Coarse  Crushing  and  Conveying. — This  is  done  in  a  sep- 
arate building.     The  ore,  broken  underground  to  9-in.  size  or  less,  is  stored 
in  two  1000-ton  flat-bottom  bins.     It  forms  a  natural  slope,  so  about  half  of 
it  can  be  drawn  off  at  the  side  gates  through  shaking  grizzlies,  having  2-in. 
openings.     The  undersize  goes  by  a  conveying  belt  6,  to  a  trommel  8,  while 
the  oversize  falls  upon  an  endless  picking  belt,  40  in.  wide,  traveling  45  ft. 
per  minute.     Here  the  ore-sorters  pick  out  the  waste  and  throws  it  down 
chutes  to  a  20-in.  conveyor  belt,  which  discharges  upon  the  waste  dump. 
The  picking  belt  discharges  upon  a  shaking  feeder  which  feeds  it  to  a 
gyratory  crusher  8.     The  crusher  discharge  joins  the  undersize  from  the 
shaking  grizzly  in  the  trommel,  4  ft.  diameter  by  14  ft.  long  that  has  1  J-in. 
openings.   Its  undersize  goes  direct  to  the  inclined  conveyor  belt  11,  and  the 
oversize  feeds  the  gyratory  crushers  set  to  crush  1-in.  size.     The  discharge 
from  the  crushers  joins  the  trommel  undersize  on  the  inclined  belt-elevator. 

Sampling. — An  inexpensive  sampling  of  the  ore  is  thus  effected.  On 
a  vertical  shaft  is  a  bevel  gear  with  a  bucket  of  5J  Ib.  capacity  attached  to 
it.  When  in  the  revolution  of  the  gear,  the  bucket  traverses  the  stream 
of  falling  ore  from  the  top  end  of  the  inclined  conveyor  it  takes  out  a 
sample.  Now,  on  the  gear,  one-third  of  the  teeth  are  removed  and  at  that 
space  a  counter-balance  actuates  the  gear,  allowing  the  bucket  to  make  a 
quick  cut  through  the  ore.  In  this  way  a  sample  of  1  ton  hi  500  is  taken 
to  be  cut  down  by  the  usual  sampling  methods. 

(2)  Stamping,  Tube-milling  and  Concentrating. — From  the  head  of 
the  inclined  belt  the  ore  is  conveyed  by  a  horizontal  belt  and  distributed 
evenly  to  flat-bottom  mill  bins  16  ft.  wide,  17  ft.  high,  by  110  ft.  long,  of 
1500   tons   capacity.      Half  of   the  ore  which  will   run   freely  is  fed  to 
the  stamp-batteries. 


266 


CYANIDATION  OF  SILVER  ORES 


Compressor  'A 

FIG.  147. — Flow-sheet  of  Belraont  Mill. 


THE  BELMONT  MILL  267 

Crushing  is  done  by  sixty  stamps  of  nearly  9  tons  duty  per  stamp 
through  4- to  6-mesh  screens.  It  is  done  in  cyanide  solution  at  the  rate 
of  5  tons  and  with  the  addition  of  1  Ib.  quicklime  per  ton  of  ore.  The 
battery  is  shown  in  Fig.  87. 

The  stamp  battery  discharge  is  delivered  to  eight  Dorr  duplex-classifiers 
placed  in  closed  circuit  with  their  respective  tube-mills  as  shown  in  Fig. 
40.  A  further  addition  of  lime  as  milk  of  lime  is  made  at  this  point,  also 
an  addition  of  0.15  Ib.  lead  acetate  per  ton  of  ore.  The  most  economical 
point  in  grinding  has  been  to  yield  a  product  75  per  cent  of  which  will  pass 
200-mesh.  The  overflow  from  the  classifiers  passes  now  to  eight  5-ft. 
Callow  cones  used  as  sloughing-off  cones;  that  is,  there  is  a  turbid  or  slimy 
overflow  going  to  Dorr  thickeners  and  a  spigot  discharge  forming  the  feed 
for  sixteen  No.  6  Wilfley  concentrating-tables  running  300  strokes  per 
minute,  with  a  f-in.  throw.  Close  concentration  is  not  attempted,  but 
rather  to  obtain  a  clean  concentrate  product.  This  is  put  into  a  vacuum 
tank  14  ft.  diameter  by  3  ft.  deep,  and  allowed  to  dry  under  vacuum  for 
forty-eight  hours.  Afterward  it  is  shoveled  from  the  tank  to  a  steel  bin 
below.  It  is  sampled  and  shipped  by  rail  to  the  smelter.  In  attempting 
to  treat  these  concentrates  on  the  spot,  it  was  found  that,  while  a  good 
extraction  could  be  made' with  fresh  solution,  soon  this  became  foul  and 
inactive  and  the  extraction  became  poor.  Even  where  this  was  good,  the 
cost  of  treatment  would  be  higher  than  the  shipping  and  marketing  expense. 
Farther  it  was  found,  even  when  roasting,  a  satisfactory  extraction  was  not 
attainable  by  cyaniding.  These  concentrates  were  about  60  per  cent 
pyrite,  30  per  cent  insoluble  and  contained  1.6  per  cent  silver  and  gold,  or 
467  oz.  per  ton.  The  tailings  pulp  from  the  Wilfley  tables  joins  the  over- 
flow from  the  Callow  cones  to  go  to  four  30  by  12-ft.  Dorr  thickeners  (see 
Fig.  102).  The  proportion  of  solution  to  ore  in  this  united  flow,  which  had 
been  added  in  the  prior  operations,  is  as  follows:  At  the  stamp  batteries 
5  to  6  parts;  to  correct  the  water  at  the  tube-mill  feed  0.8  part;  at  the 
Dorr  classifiers,  in  order  to  thin  the  pulp  for  proper  classification,  1  part; 
as  wash-water  at  the  Wilfley  tables  about  1  part.  Two  of  these  classifiers 
are  fitted  with  trays,  being  another  bottom  with  a  central  opening.  Addi- 
tional settling  area  is  thus  afforded,  increasing  their  capacity  by  about  75 
per  cent.  The  overflow  from  these,  the  first  set,  is  sent  to  the  partial 
precipitated  solution  tank  54  while  the  underflow  of  1.26  specific  gravity,  is 
pumped  to  the  first  tank  of  the  first  series  of  Pachuca  agitators  28,  and 
passes  through  them  all  in  series.  Each  of  these  tanks,  15  ft.  diameter  by 
45  ft.  high,  is  agitated  by  a  central  air-lift  (see  Fig.  100).  The  pulp  from 
the  last  agitator,  diluted  with  four  more  parts  of  partial  precipitated  solu- 
tion from  54  goes  to  a  second  set  of  Dorr  thickeners,  the  overflow  to 
the  press-solution  tank  44,  while  the  thickened  pulp  is  delivered  to  be 
treated  in  series  through  a  second  set  of  agitators,  whereby  the  silver  has 


268  CYANIDATION  OF  SILVER  ORES 

been  thoroughly  brought  into  solution.  The  flow  from  the  last  agitator 
goes  to  33,  a  filter  stock-tank  28  diameter  by  20  ft.  high,  equipped  with  a 
Trent  mechanical  agitator  to  keep  the  pulp  from  settling.  The  total  period 
for  agitation  is  forty-eight  hours.  By  thus  thinning  the  pulp  with  a 
nearly  barren  solution  after  it  has  passed  through  the  first  series  of  agitators 
a  large  proportion  of  the  dissolved  silver  is  displaced,  which  relieves  the 
vacuum  filters  36,  from  much  work,  and  allows  a  change  of  solution  for 
final  agitation.  At  No.  1  agitator  enough  cyanide  has  been  added  to  bring 
up  the  solution-strength  to  6  Ib.  cyanide  per  ton;  also  0.18  Ib.  per  ton 
of  lead  acetate  is  then  put  in.  In  the  central  columns  of  the  agitators  are 
steam  coils  for  heating  the  rapidly  rising  slime,  it  having  been  found  that 
by  thus  heating  a  2  per  cent  better  extraction  is  attained.  The  watery 
slime  in  the  filter  stock-tank  is  now  ready  for  filtering.  This  is  accom- 
plished at  the  filter  boxes  or  tanks,  of  250  filter  leaves.  The  pregnant 
solution  from  the  vacuum  filter  boxes  (now  to  be  precipitated)  is  pumped 
to  the  press  solution  tank  44,  which  is  30  ft.  diameter  by  10  ft.  deep.  It  is 
treated  in  a  Crowe  vacuum  cylinder,  clarified  and  sent  on  to  the  partial 
precipitation  or  complete-precipitation  supply  tanks. 

The  Refinery. — Here,  for  partial  precipitation  are  three  Merrill  trian- 
gular presses  and  for  complete  precipitation  one  press.  Only  enough  solu- 
tion is  completely  precipitated  to  a  value  of  about  10  cents  per  ton,  and  is 
used  for  dilution  at  the  second  set  of  Dorr  thickeners  for  table-wash  solu- 
tion and  tube-mill  feed-solution.  The  completely  precipitated  product 
carries  75  per  cent  silver  (and  gold),  7  per  cent  insoluble  matter,  3.6  per 
cent  lime  carbonate,  and  nearly  5  per  cent  zinc  and  other  impurities, 
notably  selenium  to  the  extent  of  J  per  cent.  For  a  clean-up  the  product 
is  well  mixed  with  2^  per  cent  of  borax  and  6  per  cent  each  of  soda  and 
sand,  and  made  into  briquettes.  These  briquettes  are  melted  in  double- 
compartment  carborundum-lined  furnaces,  Fig.  146,  to  produce  a  bullion 
containing  93  per  cent  silver  (and  gold),  2.5  per  cent  lead  and  1.8  per  cent 
selenium. 

On  about  half  a  million  tons  of  ore  the  extraction  of  silver  was  by  con- 
centration 9  per  cent;  by  cyanidation,  84  per  cent,  a  total  of  93  per  cent. 
Of  the  gold  6  per  cent  was  recovered  by  concentration,  90.2  per  cent  by 
cyaniding. 

CYANIDATION  OF  MIXED   SILVER  ORES  AT  THE   SAN  FRANCISCO   MILL, 

PACHUCA,  MEX. 

This  plant  treats  ores  from  a  number  of  different  mines,  each  of 
which  is  different  in  chemical  composition  and  requires  a  variation  in  the 
treatment.  The  first  Pachuca  tanks  used  in  North  America  were  installed 
at  this  plant. 


THE  SAN  FRANCISCO  MILL  269 

The  average  content  of  the  ore  received  has  been  17.4  oz.  silver,  and 
0.08  oz.  gold  per  ton. 

The  plant  contains  a  complete  sampling  mill  where  all  ores  are  received, 
crushed,  automatically  sampled,  and  dumped  into  bins,  from  which  they 
are  carried  to  the  battery-bins  in  cars.  The  batteries  consist  of  forty 
stamps,  weighing  1250  Ib.  each,  which  drop  7J  in.  at  the  rate  of  100  drops 
per  minute.  The  ore  is  crushed  in  the  batteries  to  pass  through  a  No.  9 
roll  slot-screen  and  thence  falls  onto  eight  Wilfley  concentrating  tables, 
whence  the  tailing  runs  into  four  Dorr  classifiers,  from  which  the  slime  and 
excess  solution  fall  into  four  Frenier  pumps,  by  which  they  are  elevated  to 
two  Dorr  slime-thickeners,  24  ft.  in  diameter  by  10  ft.  deep.  The  sand 
from  the  Dorr  classifiers  falls  into  four  tube-mills,  4J  ft.  in  diameter  by 
13  ft.  long,  so  placed  that  each  mill  receives  the  sand  from  one  classifier. 
After  the  sand  is  reground  in  the  tube-mills,  it  passes  into  four  Frenier 
pumps,  which  return  it  to  the  Dorr  classifiers  until  the  whole  of  the  pulp 
has  been  converted  into  slime  which  passes  to  the  two  Dorr  slime-thickeners 
previously  mentioned. 

From  the  bottom  of  these  slime-thickeners  the  slime  of  the  proper 
consistence  for  good  concentration,  is  drawn  off  and  fed  to  fifteen  concen- 
trating tables  by  means  of  which  the  heavy  minerals  which  have  been  lib- 
erated by  regrinding,  or  which  escaped  concentration  on  the  Wilfleys,  are 
concentrated  out  of  the  slime.  The  tailings  from  the  concentrating  tables 
are  elevated  by  a  3-in.  centrifugal  pump  to  a  third  Dorr  slime-thickener 
24  ft.  in  diameter  by  10  deep,  whence  the  thickened  slime,  containing  1  ton 
of  dry  slime  to  1  \  of  solution,  flows  to  the  Pachuca  vats  for  agitation,  while 
the  overflow  of  this,  as  well  as  the  overflow  of  the  other  two  slime-thick- 
eners, flows  to  a  tank  from  which,  when  clear,  it  is  pumped  to  the  vats 
above  the  mill  which  supply  solution  to  the  batteries.  It  sometimes  occurs, 
when  milling  certain  classes  of  ore,  that  the  third  Dorr  slime-thickener  will 
not  have  a  sufficient  capacity  for  settlement  of  the  ore  milled,  so  that  the 
overflow  contains  unsettled  slime.  When  this  is  the  case  the  overflow 
from  this  thickener  is  run  into  a  series  of  four  masonry  settling- tanks, 
where  the  slime  is  settled  before  pumping  the  solution  to  the  vats  which 
supply  the  batteries. 

There  are  eight  Pachuca  tanks  in  this  plant,  each  of  which  is  charged 
with  from  80  to  100  tons  of  dry  slime  for  treatment.  The  slime  resulting 
from  the  milling  and  classification,  which  is  treated  in  these  tanks,  is  of 
such  a  fineness  that  80  per  cent  will  pass  through  a  200-mesh  screen. 

The  slime  is  elevated  by  pumping  to  a  slime  storage  tank  placed  above 
the  filters,  from  which  it  is  fed  to  the  filters  as  required.  There  are  two 
filter-plants  in  this  mill,  each  having  a  capacity  of  150  tons  per  day.  The 
Butters  filter  has  seventy-eight  filter-leaves,  while  the  Moore  filter  has  two 
baskets  of  forty  leaves  each.  Each  filter-plant  is  supplied  with  all  of  the 


270  CYANIDATION  OF  SILVER  ORES 

latest  improvements  and  they  give  entire  satisfaction.  After  filtration 
the  solutions  are  pumped  to  the  sand  niters  above  the  precipitation  room 
while  the  slime  tailing,  after  being  filtered  and  washed,  is  discharged  into 
the  tailing-dam.  The  solution  from  the  tailing-dam,  after  settlement  of 
the  slime,  is  pumped  to  a  tank  above  the  filters  for  use  .in  them  as  wash- 
water. 

The  precipitation  room  contains  ten  zinc-boxes  made  of  sheet  steel, 
each  15  ft.  long,  3  ft.  wide,  and  2|  ft.  deep  which  are  divided  into  five 
compartments  each.  The  precipitation  of  the  clarified  solutions  from  the 
sand  filters  is  almost  perfect  in  these  boxes,  as  the  solutions  entering  the 
boxes  assay  from  100  to  250  gm.  of  silver  and  from  1  to  3  gm.  of  gold  per 
ton,  while  the  precipitated  solutions  leaving  the  boxes  do  not  carry  more 
than  1  gm.  of  silver  and  0.1  gm  of  gold  per  ton. 

The  cakes  of  precipitate  from  the  filter-press  are  dried  until  they  con- 
tain from  5  to  10  per  cent  of  moisture.  Then,  after  breaking  up  the  larger 
lumps,  a  mixture  is  made,  consisting  of  precipitate  81  per  cent,  sodium 
carbonate  7  per  cent,  borax  glass  10  per  cent,  quartz  tailing,  2  per  cent. 

A  graphite  crucible,  No.  300,  is  placed  in  the  coke  furnace,  and,  as  soon 
as  it  begins  to  heat  up,  is  filled  with  the  mixture  of  precipitate  and  flux, 
and  melted  in  the  usual  way.  Total  treatment  costs  are  $1.93  per  ton, 
with  recovery  of  92.4  per  cent  of  the  silver  and  94.5  per  cent  of  the  gold. 

THE  WAIHI  GRAND  JUNCTION  MILL,  WAIHI,  NEW  ZEALAND 

The  ore  from  the  mine  consists  essentially  of  a  gangue  of  quartz  and 
calcite,  with  8  to  10  per  cent  sulphide — pyrite,  sphalerite,  galena,  chalco- 
pyrite,  and  traces  of  arsenic  and  antimony.  In  stoping,  this  is  mined  with 
the  well-defined  lode,  thus  bringing  into  the  mill  at  times  a  considerable 
quantity  of  hard  country  rock.  The  gold  exists  in  an  exceedingly  fine 
state  of  division,  consequently  rendering  it  necessary  to  grind  very  fine  to 
obtain  a  satisfactory  extraction.  The  ore  averages  about  $8  gold  and  $1 
silver  per  ton.  There  are  forty  1100-lb.  stamps  fed  by  a  hanging  type  of 
improved  Challenge  ore-feeder.  Cyanide  solution  of  0.10  per  cent  is  added 
here  in  the  proportion  of  10  of  solution  to  one  of  ore,  also  lead  acetate  equal 
to  0.5  Ib.  per  ton  of  ore. 

From  the  stamps  with  a  duty  of  7.6  tons  a  day  the  pulp  flows  to  three 
elevator  wheels,1  by  means  of  which  it  is  raised  to  a  series  of  conical  boxes 
or  spitzkasten.  The  overflow  from  these  boxes  passes  on  to  the  Wilfley 
tables.  The  underflow,  after  passing  through  the  tube-mills,  is  again 
returned  by  the  same  elevators  to  the  conical  boxes.  The  Wilfley  tables 
are  used  merely  as  classifiers.  The  discharge  is  taken  from  18  in.  along  the 

1  The  tailings  wheel,  though  of  greater  first  cost,  is  reliable  and  the  cost  of  repairs  is 
small. 


THE  WAIHI  GRAND  JUNCTION  MILL 


271 


side,  and  the  tables  are  so  set  that  their  overflow  is  practically  fine  enough 
for  economic  treatment.    The  heads,  consisting  of  the  larger  particles  of 


_/B^\£^  Sampler 1  *$»*$£ 


^-^*V_  Strong  Solution  Clarifiers 

(-EC_<__ir^ — EL.-^jri — JT    \ 


Water 
.Cyanide  Solution 


FIG.  148. — Flow-steet  of  Waihi  Grand  Junction  Co.'s  Mill. 


concentrate  and  sand,  are  sluiced  to  belt  elevators,  which  raise  and  dis- 
charge into  the  feed-cones  of  the  tube-mills. 


272  CYANIDATION  OF  SILVER  ORES 

There  are  fourteen  tube-mills,  run  at  27  R.P.M.  The  average  life  of 
the  liners  is  eighteen  months.  The  angle  shoes  are  also  made  of. hard 
cast  iron,  and  have  an  average  life  of  ten  months.  The  pulp  is  fed  into  the 
mills  by  means  of  injectors.  The  feed  nozzles  are  1J  in.  inside  diameter 
and  discharge  into  larger  cast-iron  pipes  which  are  bolted  on  to  the  end  of 
the  mill.  The  clearance  between  these  pipes  is  |  in.  The  ratio  of  solution 
to  ore  in  feed  is  as  1  is  to  1.  Cost  of  tube-milling  is  about  28  cents  per  ton. 

The  overflow  of  the  Wilfley  tables  flows  to  a  spitzlutte,  the  spigot 
product  of  which  is  returned  to  the  tube-mills  by  a  centrifugal  pump.  The 
overflow,  consisting  of  solution  and  slime  in  the  proportion  of  10  to  1, 
flows  to  the  settlers  or  thickeners,  lime  being  added  to  the  launder  that 
conveys  it.  Of  the  overflow  of  clear  solution,  from  the  settlers,  75  per  cent 
flows  to  the  strong  solution  sump  and  25  per  cent  to  the  strong  solution 
clarifying  tanks.  The  pulp  is  drawn  continuously  from  the  bottom  of  the 
settlers,  and  is  of  a  consistence  of  1^  of  solution  to  1  of  slime.  This  is 
pumped  into  one  of  a  series  of  12  flat  agitators,  each  22|  ft.  by  6  ft.,  pro- 
vided with  four  arms  revolving  at  the  rate  of  4  R.P.M.  These  agitators 
are  used  as  storage  tanks.  When  three  are  full  they  are  discharged  by  a 
pump  into  a  tall  tank,  55  ft.  by  13J  ft.  Agitation  is  by  compressed  air 
at  a  pressure  of  38  Ib.  per  square  inch.  Each  tank  has  a  capacity  of  250 
tons  of  pulp.  Agitation  is  continued  in  these  tanks  for  an  average  of 
18  hours  and  gives  a  further  extraction  of  12  cents  per  ton.  The  strength 
of  cyanide  is  maintained  by  a  continuous  flow  of  strong  stock  solution, 
which  is  added  to  that  supplying  the  mortar  boxes.  This  is  the  only  place 
where  cyanide  is  added.  The  consumption  of  KCN  is  1.258  Ib.  per  ton. 

From  the  tall  tanks  the  pulp  gravitates  to  the  Moore  filter  plant. 
Each  basket  is  composed  of  ten  frames,  16  ft.  by  4  ft.,  giving  a  total  filter- 
ing area  of  1280  sq.  ft.  per  basket.  The  weight  of  slime  cakes  in  each 
basket  is  equivalent  to  five  tons  dry  weight;  seventy-five  minutes  is 
required  for  formation  with  a  vacuum  of  25  in.  of  mercury.  The  loading 
tanks  are  fitted  with  conical  bottoms,  and  an  air-lift  is  arranged  to  prevent 
settlement  of  slime.  The  basket  is  raised  by  an  overhead  traveling  crane, 
electrically  driven,  and  transferred  first  to  a  wash-tank  of  weak  solution 
(0.04  per  cent  KCN),  where  the  charge  is  washed  for  forty  minutes,  then 
to  a  water  tank,  where  a  further  wash  of  water  is  drawn  through  for  ten 
minutes.  The  basket  is  now  raised  and  suspended  over  the  discharge 
tank,  drained  thoroughly,  and  the  tailing  sampled  for  assay.  When  the 
vacuum  is  destroyed  the  charge  falls  on  to  the  revolving  arms  in  the  tank, 
and  is  sluiced  away  with  water  jets  to  the  sludge  channel. 

The  cost  of  treatment,  which  includes  agitation  and  vacuum  filtration, 
is  64  cents  per  ton. 

The  vats,  receiving  the  solution  from  the  slime  treatment,  are  provided 
with  filters  to  prevent  slime  from  passing  into  the  zinc-boxes.  From  these 


THE  WAIHI  GRAND  JUNCTION  MILL  273 

vats  the  gold-bearing  solutions  flow  through  meters,  which  register  the 
tonnage,  to  the  launder  feeding  the  extractor  boxes.  There  are  twenty-two 
boxes,  seventeen  of  which  are  used  for  strong  solution  (0.10  per  cent  KCN) 
and  five  for  weak  solution  (0.0  4  per  cent  KCN).  The  strong  solution  boxes 
have  eight  compartments,  with  total  available  zinc  space  of  740  cu.  ft.; 
the  weak  solution  boxes  have  a  total  zinc  space  of  145  cu.  ft.  All  are  pro- 
vided with  a  screen  of  4-mesh  near  the  bottom  of  each  compartment.  The 
average  rate  of  flow  of  solution  through  the  strong  boxes  is  1.4  tons  per 
cubic  foot  of  zinc  per  twenty-four  hours,  and,  through  the  weak,  1.1  tons. 
Precipitation  of  the  metals  from  the  weak  solution  is  assisted  by  first  dip- 
ping the  zinc  in  a  solution  of  lead  acetate.  At  the  head  and  tail  of  each  set 
of  boxes  is  an  automatic  solution  sampler.  These  samples  are  assayed 
daily,  and  from  the  results,  together  with  the  tonnage  passed  through  the 
boxes  as  registered  by  the  meters,  the  amount  of  bullion  precipitated  can  be 
calculated.  The  bulk  of  the  precipitation  takes  place  in  the  first  two  com- 
partments. The  first  four  are  cleaned  out  every  seven  days,  and  all  the 
compartments  every  twenty-eight  days — that  is,  at  the  end  of  each  period. 
When  the  boxes  are  cleaned  the  zinc  is  washed  in  the  first  cells,  the  pre- 
cipitate is  drawn  off  by  means  of  a  suction  hose  into  a  receiving  cylinder, 
thence  into  vacuum  filters.  The  short  zinc,  which  passes  4-mesh  and 
remains  on  25-mesh,  is  treated  with  1  to  6  per  cent  sulphuric  acid  in  vats 
fitted  with  revolving  arms,  which  agitate  the  charge,  which,  after  being  well 
washed  with  hot  water,  is  dried  with  the  balance  of  the  precipitate.  The 
precipitate  is  dried  in  cast-iron  ovens  without  stirring.  The  slime  and 
fluxes  are  mixed  in  a  small  tube-mill,  which  is  entirely  enclosed  and  dust- 
proof,  and  when  thoroughly  mixed  it  is  discharged  into  a  truck.  Melting 
of  the  precipitate  is  conducted  in  No.  120  graphite  crucibles.  Kerosene 
fuel  is  used,  and,  with  four  furnaces,  two  men  melt  and  refine  1 100  Ib.  of 
precipitate  in  eight  hours.  The  bullion  and  slag  are  poured  together  into 
conical  molds,  and,  when  solid,  the  bullion  is  detached  and  re-melted  in 
the  same  crucibles  and  poured  into  bars  weighing  about  1100  oz.  each. 
The  slag  is  crushed  by  stampers  through  25-mesh  screen,  passed  over  blan- 
keting to  retain  prills  of  bullion,  and  collected  in  settling  boxes.  It  is 
afterward  air-dried,  bagged,  and  shipped  to  smelters  for  treatment.  The 
cost  of  precipitation  and  melting  is  5  cents  per  ton  of  ore  crushed  in 
mill. 

Two  Balback  tilting  furnaces,  using  coke  fuel,  have  been  installed,  and 
are  used  for  melting  the  precipitate,  the  kerosene  furnaces  now  being  used 
for  remelting  the  bullion  into  bars. 

The  total  cost  of  milling  and  treatment  is  $1.50  per  ton. 


274  CYANIDATION  OF  SILVER  ORES 

MILLING  PRACTICE  AT  COBALT,  ONTARIO 

There  are  two  grades  of  ore,  the  high  and  low  grades,  each  treated  by  a 
different  process.  The  treatment  of  the  high-grade  ore  is  given  on  page 
243 ;  the  milling  of  the  low  grade  is  described  below. 

The  Ore. — This  is  complex,  containing  chiefly  native  silver  occurring  in 
particles  entangled  in  a  mixture  of  nickel  and  arsenic  minerals  (smaltite 
and  niccolite)  and  calcite.  It  is  classified  in  two  grades,  the  "  high  grade," 
chiefly  of  native  silver  of  2500  oz.  per  ton,  and  the  "  low-grade,"  having 
native  silver,  also  considerable  sulphides  and  antimonides,  whose  silver  is 
difficult  to  recover. 

Preliminary  Treatment. — At  the  washing  plant,  as  shown  on  the  flow 
•sheet,  Fig.  149,  the  mine  ore  is  delivered  to  a  trommel  40  in.  diameter  by 
10  ft.  long,  where  with  the  addition  of  water  it  is  washed  and  separated 
into  two  products,  the  oversize  going  to  two  sets  of  trommels  to  yield 
products  for  two  sets  of  jigs.  The  jigs  yield  a  high-grade  head  which 
joins  the  corresponding  product  from  the  30-in.  picking  belt  to  go  to  the 
high-grade  ore  mill.  This  ore  amounts  to  about  9  per  cent  of  the  total,  and 
its  treatment  in  the  high-grade  mill,  as  indicated  at  the  right  side  of  the 
flow-sheet,  is  described  under  head  of  "  amalgamation  of  silver  ores," 
page  243.  This  preliminary  treatment  lessens  the  period  of  agitation 
necessary  with  cyanide  solution. 

Many  experiments  involving  the  preliminary  reduction  of  telluride 
gold  ores  in  caustic  soda  solution  with  aluminum  and  zinc,  as  well  as 
cathodic  reduction  in  both  caustic  soda  and  salt  solution,  have  shown  that 
with  cheap  fuel  and  efficient  roasting  this  form  of  preliminary  treatment 
would  present  no  advantage.  The  gold  telluride  compounds  are  in  general 
more  difficult  to  reduce  than  the  sulphide  and  sulpho-antimonide  silver 
minerals. 

To  go  back:  the  undersize  from  the  last  trommel,  together  with  the  jig 
tailings,  is  elevated  to  a  dewatering  trommel,  30  in.  diameter  by  6  ft.  long. 
The  oversize  joins  the  ore  from  the  picking  belt  destined  for  the  low-grade 
mill,  while  the  undersize  after  dewatering  is  sent  to  join  the  battery  pulp 
in  the  same  mill. 

THE  NIPISSING  CO.'S  "  LOW-GRADE "   MILL,   COBALT,  ONTARIO    i 

The  ore,  of  about  26  oz.  silver  per  ton  is  treated  by  all  sliming  cyaniding. 
The  left-hand  portion  of  Fig.  149  gives  the  flow-sheet  of  this  mill. 
There  are  forty  heavy  stamps  of  1400  lb.,  each  battery  of  ten  stamps 
being  driven  by  a  40  H.P.  motor,  wet-crushing  through  a  2-  to  3-mesh 
screen  and  using  for  each  ton  crushed  7  tons  of  a  solution  containing  0.7  lb. 
of  caustic  soda  per  ton.  This  caustic  soda  addition  is  necessary  as  a  pre- 
liminary to  the  desulphurizing  treatment  to  follow. 


THE  NIPISSiNG  LOW-GRADE  MILL 


275 


-j-  Ore  Bins 
4-  Scales 
4-  Washing  Trommel 

Crush 

1 

• 

! 

ing  Solution      : 

| 

To  Low  pra 

•  Screening  Trommels 
Washing  Plant 

-  .  Pi«king  Belt 

:  2  Sets  Jigs 
s~*\^~*~~^~" 

•Sor^D^BtrTJewaterer 

-  Gyratory 
Crusher 

de  Ore  Mill                                    To  High  grade  Ore  Mill 

_^ 

-Battery 
Tube  Mills  and  Classifiers 

•  Cms 
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Cyanide- 
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Amalgam  Bags 
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Precipitates 
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Solution  Tat 

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Discharged 
for  Coba 

FIG.  149. — Flow-sheet  of  Nipissing  Co.'s  Mill. 


276  CYANIDATION  OF  SILVER  ORES 

Classification  and  Fine  Grinding. — For  high  extraction  the  ore  must 
all  be  ground  to  pass  200-mesh.  The  battery  pulp  is  classified  in  a  first 
set  of  four  Dorr  classifiers,  the  slime  overflowing  direct  to  one  or  other  of 
the  slime-collecting  vats  in  the  cyanide  plant,  the  oversize  going  to  two 
coarse-grinding  tube-mills.  The  discharge  from  these  mills  flows  into  a 
second  set  of  Dorr  classifiers  whose  overflow  joins  that  of  the  first  set  at 
the  slime-collecting  vats,  and  the  sands  are  elevated  to  two  fine-grinding 
tube  mills.  The  discharge  from  these  flow  into  a  drag  classifier,  then  back 
to  the  second  set  of  Dorr  classifiers  all  in  closed  circuit  with  the  fine-grinding 
mills. 

Cyanide  Plant. — This  is  in  the  largest  building,  where  the  vats  are 
placed  in  two  rows,  and,  with  the  exception  of  the  solution  vat,  fitted  with 
mechanical  stirrers  driven  by  a  125  H.P.  motor  at  8  R.P.M.  The  working 
load  in  each  vat  is  140  tons  of  dry  slime  with  280  tons  of  solution.  The 
slime-collecting  vats  are  dewaterers  delivering  a  clear  overflow  to  a  lower 
crushing  solution  vat  and  thence  by  pipe  line  marked  "  caustic  soda  solu- 
tion," to  the  upper  crushing  vat  for  re-use  at  the  battery.  When  one  col- 
lecting vat  is  full  of  a  charge  the  pulp  flow  is  switched  to  the  other.  The 
charge  is  then  agitated  for  an  hour  and  the  thickened  pulp,  consisting  of  1.5 
parts  of  caustic  soda  solution  to  one  of  the  pulp,  is  pumped  to  desul- 
phurizing treatment. 

The  Wet  Desulphurizing  Process. — This  breaks  up  the  refractory 
silver  minerals,  the  slimed  pulp  being  brought  in  contact  with  aluminum 
in  the  caustic  soda  solution  in  a  tube-mill  and  the  silver  left  in  the  spongy 
metallic  state  amenable  to  cyanide  treatment.  To  do  this  the  collecting 
vat  pulp  passes  to  a  desulphurizing  tube  mill,  revolving  10  R.P.M. 
where  it  comes  in  contact  with  a  load  of  about  4000  Ib.  of  aluminum  pieces 
of  1J-  to  2-in.  cubes.  The  pulp  discharge  from  the  mill  gravitates  to  the 
desulphurizing  vat  adjoining.  In  this  tank  34  ft.  diameter  by  13  ft.  deep 
and  lined  with  aluminum  in  plates,  the  pulp  receives  mechanical  agitation 
for  a  period  of  twenty-four  to  thirty-six  hours.  In  order  to  keep  the  mill 
cyanide  solution  in  balance  it  is  necessary  to  eliminate  as  much  as  possible 
of  the  caustic  soda  solution  from  the  desulphurized  pulp.  This  is  done  in  a 
60-leaf  Butters  vacuum  filter,  marked  "  alkaline  filter  "  in  the  plan  view. 
This  preliminary  treatment  lessens  the  period  of  agitation  necessary  with 
cyanide  solutions. 

The  filtrate  passes  to  the  crushing  solution  vat  while  the  thick  slime  is 
pumped  to  one  of  the  seven  cyanide  treatment  vats  fitted  for  mechanical 
agitation.  There  it  is  treated  in  charges  of  130  tons  of  dry  slime  to  260 
tons  of  cyanide  solution  of  0.25  per  cent  (5  Ib.  per  ton)  with  0.2  per  cent 
alkali.  An  air  lift  is  operated  constantly  during  agitation,  the  pulp  being 
drawn  off  from  the  bottom  and  discharged  into  the  top  of  the  vat  and 
treated  at  the  same  time.  After  agitation  the  charge  is  allowed  to  settle 


THE  NIPISSING  LOW-GRADE  MILL 


277 


so  that  the  clear  solution  may  be  decanted  to  the  "  pregnant  solution  vat." 
The  pulp  is  then  stirred  and  pumped  to  a  34-ft.  cyanide  stock-pulp  vat, 
where  it  must  be  kept  agitated  until  drawn  off  to  the  80-leaf  Butters  vacuum 
"  cyanide  filter."  The  clear  solution  from  the  filter  is  delivered  to  the 
"  pregnant  solution  "  tank,  the  residue  slime  is  discharged  to  the  residue 
dump. 

The  pregnant  solution  is  now  ready  to  go  to  the  precipitation  depart- 
ment at  the  right  of  the  tube-mill  house  to  be  subjected  to  aluminum  pre- 
cipitation treatment  for  twenty-four  hours. 

The  cost  of  treatment  in  1912  was  given  as  $3  per  ton.  The  cost  of  the 
plant  $254,000  to  treat  244  tons  daily. 

CYANIDATION  OF  SILVER-BEARING  CONCENTRATES 


C 

ASSAY  oz. 

PER  TON. 

%  EX- 
TRACTION. 

COSTS  PER  TON  OF  CONCENTRATES. 

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20.0 

6  25 

C.  Tube-milled  in  1  Ib.  KCN   solution;    agitated  in  6  Ib.  KCN  for 
ten  to  fourteen  days,  with  two  decantations;   filter  pressed  and  precip- 
itated by  zinc  dust;  2  Ib.  lead  acetate  and  1  Ib.  mercuric  chloride  per  ton 
of  concentrates. 

D.  Tube-milled  and  agitated  in  3  Ib.  KCN  solution  for  eight  to  ten 
days. 


CHAPTER   XXIV 
PARTING  SILVER-GOLD  BULLION 

PARTING  SILVER-GOLD  INGOTS  OR  BARS  WITH  ACIDS 

The  bars  from  reduction  works  commonly  contain  gold  and  silver 
alloyed  with  copper,  but  sometimes  also  zinc  and  lead.  It  is  customary 
to  remelt  the  bars  and  assay  them,  buying  them  on  the  result  of  the  assay. 

The  bars  are  parted  in  nitric  or  sulphuric  acid.  Sulphuric  acid,  being 
cheaper,  is  the  acid  commonly  employed. 

Bars  containing  a  large  proportion  of  gold  are  inquartated  by  melting 
them  with  silver  in  order  to  decrease  the  ratio  of  gold  to  silver  to  at  least 
1  to  2.5,  otherwise  the  acid  fails  to  attack  the  silver.  In  parting  with  sul- 
phuric acid  the  copper  should  be  less  than  10  per  cent,  but  in  nitric-acid 
parting  more  than  10  per  cent  is  allowed.  To  adjust  this  percentage,  bars 
low  in  copper  are  melted  with  those  high  in  that  metal. 

Nitric-acid  Parting. — This  method,  still  practiced  at  the  United  States 
Mint,  Philadelphia,  is  an  efficient  way  of  parting,  especially  on  a  small  scale. 
The  bars  having  been  melted  and  proportioned  as  above  described,  the 
molten  metal  is  granulated  by  pouring  it  into  a  tank  of  water.  The  gran- 
ulated metal  is  transferred  to  porcelain,  glass,  or  platinum  vessels,  and 
treated  with  nitric  acid,  1.20  sp.  gr.,  until  action  ceases.  The  solution  is 
allowed  to  settle,  then  is  decanted,  and  fresh  acid  is  added  for  the  purpose  of 
dissolving  the  remaining  traces  of  silver.  This  is  again  decanted,  and  the 
gold  residue  is  washed  thoroughly  with  hot  water.  It  is  then  ladled  out, 
drained,  dried,  and  mixed  with  a  little  flux,  and  melted  in  a  graphite  cru- 
cible. 

From  the  decanted  silver  solution  the  metal  is  precipitated  by  adding 
common  salt.  The  silver  chloride  thus  formed  is  washed,  thoroughly 
granulated,  zinc  is  added  to  reduce  the  silver  to  metal,  and  the  zinc  chloride 
resulting  from  the  reaction  is  washed  from  the  precipitated  silver  A  The 
silver  is  pressed  into  cakes,  melted,  and  cast  into  ingots  for  bar-silver. 

Sulphuric-acid  Parting. — Since  for  this  method  of  parting  there  should 
be  less  than  10  per  cent  copper  present  in  the  gold-silver  alloy,  and  not  less 
than  2.5  parts  silver  to  1  of  gold,  the  bars  to  be  parted  that  exceed  the 
required  limit,  are  so  selected  and  melted  with  others  as  to  afford  the 
required  proportion.  The  melted  metal  is  cast  into  flat  ingots  and  parted 
in  this  form. 

278 


PARTING  SILVER-GOLD  BULLION  279 

The  ingots  are  placed  in  a  cast-iron  kettle,  covered  with  a  sheet-iron 
hood  that  is  connected  with  a  chimney  so  that  the  acid  fumes  from  the  kettle 
are  carried  away.  Here  they  are  treated  with  sulphuric  acid  of  full  strength 
(66°  Be*.)-  When  action  has  ceased  the  solution  is  allowed  to  settle,  after 
which  the  clear  supernatant  part  is  decanted,  being  drawn  off  by  a  lead- 
pipe  siphon  into  a  lead-lined  precipitating  tank.  The  residue  in  the  kettle 
is  treated  six  or  seven  times  with  fresh  boiling  acid.  In  this  way  the  silver 
completely  dissolves,  the  acid  solution  being  removed  after  each  treatment. 
The  brown  gold  residue  is  finally  boiled  with  water,  being  heated  and 
agitated  by  live  steam  from  a  pipe  inserted  in  the  water.  In  this  way  the 
gold  is  "  sweetened."  The  residue  is  removed  from  the  kettle,  dried, 
melted  in  crucible  with  a  little  borax  for  flux,  and  cast  into  a  bar  of  gold 
999  fine. 

The  acid  solution  from  the  kettles,  which  flows  to  the  precipitating 
tank,  is  diluted  with  water,  and  the  silver  is  precipitated  by  hanging  copper 
plates  about  1  in.  in  thickness  in  the  solution.  The  copper  replaces  the 
silver  in  the  acid  solution,  which  becomes  blue  in  color.  When  precipita- 
tion is  complete  the  clear  solution  is  decanted,  and  the  cement  silver  at  the 
bottom  of  the  tank  is  washed  with  hot  water  to  remove  the  acid  copper- 
solution.  The  "  cement "  silver,  or  precipitated  silver,  is  removed  to  a 
box,  then  pressed  into  cakes  or  cheeses  in  a  hydraulic  press.  Thus  com- 
pressed, it  is  ready  for  melting  in  plumbago  crucibles  after  adding  a  little 
borax  flux.  In  large  establishments  the  silver  is  melted  in  a  small  rever- 
beratory  furnace  where  it  can  be  conveniently  fluxed,  skimmed,  and  ladled 
into  bars  for  the  market.  The  bars  weigh  35  Ib.  or  500  oz.  each.  On 
refining,  each  bar  is  marked  with  a  number  and  the  exact  weight  and 
fineness,  and  the  name  of  the  refinery  that  produced  it. 

ELECTROLYTIC  PARTING  OF  GOLD  FROM  SILVER 

The  U.  S.  Mint,  San  Francisco,  receives  bullion  of  200  fine  or  over  of 
gold  and  silver.  Gold  bars  of  900  fine  or  over  are  treated  by  a  gold-refining 
process,  while  others  are  melted  together  to  be  cast  into  cathodes  600  fine 
in  silver,  300  in  gold,  and  the  remaining  100  in  base  metals. 

Silver  Refining. — The  anodes,  3J  in.  wide  by  8|  in.  high  by  i  in.  thick, 
are  hung  in  earthenware  cells  together  with  thin  sheets  of  silver  for  cathodes, 
using  for  an  electrolyte  a  nitrate  of  silver  solution  that  carries  3  per  cent 
silver  and  1J  to  2^  per  cent  of  free  nitric  acid  and  using  a  current  of  0.8 
ampere  per  square  foot.  The  pure  silver  collects  on  the  cathodes  as  a 
sponge  deposit.  Part  falls  to  the  bottom  of  the  cell  as  slime.  From  time  to 
time  the  anodes  are  taken  out  and  the  sponge  jarred  off.  This  and  the 
slime  are  melted  to  produce  a  gold  bar.  The  pure  silver  collects  on  the 
cathodes  in  crystalline  condition  and  they  are  lifted  out  daily  to  remove 
the  silver.  The  final  product  is  melted  into  silver  bars. 


280  PARTING  SILVER-GOLD  BULLION 

Gold  Refining. — The  anodes  of  the  same  size  as  the  silver  ones  should  be 
at  least  900  fine  in  gold  and  less  than  70  fine  in  silver.  They  are  hung  in 
porcelain  cells  with  thin  pure  gold  sheets  for  cathodes  and  in  an  electrolyte 
consisting  of  gold  trichloride  having  7  per  cent  gold  and  10  to  12  per  cent 
of  free  hydrochlorine  acid.  A  high  current  (90  amperes  per  square  foot) 
is  used.  The  cathodes  when  built  up  to  160  oz.  each  are  removed  and  cast 
into  ingots  of  999  fine. 

The  spent  electrolyte  is  replaced  by  fresh.  When  it  gets  impure  it  is 
separately  treated  to  recover  some  silver  and  the  base  metals. 

PRICES  AND  COSTS 

The  prices  of  silver  ores  sold  to  custom  smelteries  are  given  under  head 
of  the  schedule  of  silver-lead  ores  being  for  a  dry  ore. 

Silver. — At  New  York  the  quotations  are  on  silver  bars,  per  troy  ounce 
of  silver,  1000  fine.  It  takes  14.58  troy  ounces  to  make  1  Ib.  avoirdupois. 
London  prices  are  for  sterling  silver,  925  fine.  The  value  of  the  pound 
sterling  is  also  given,  so  that  with  the  London  quotation,  we  may  compute 
the  equivalent  price  in  cents  there.  Let  us  say  that  sterling  exchange  is 
$4.86,  and  that  silver  is  selling  at  25d.  per  sterling  ounce,  we  have  then: 

25  1000 

TT  X  4.86  X-oF  =  54.8  cents  per  ounce  of  fine  silver. 


PART  IV 
IRON  AND  STEEL 


CHAPTER  XXV 
IRON    ORES    AND    THEIR    SMELTING 

CLASSIFICATION  AND  OCCURRENCE  OF  IRON  ORES 

These,  the  oxides  and  carbonates  of  iron  occur  accompanied  by  earthy 
minerals  forming  the  "  gangue."  Only  those  are  regarded  as  iron  ores  that 
contain  sufficient  iron  to  make  the  recovery  of  the  metal  profitable.  In 
this  discussion  of  iron  ore  it  is  to  be  understood  that  reference  is  made  to 
them  as  smelted  for  the  iron  they  contain,  and  not  as  a  flux  as  used  in 
silver-lead  and  copper  smelting  where  the  iron  enters  a  slag  to  be  thrown 
away. 

There  is  a  general  division  of  iron  ores  into  two  classes,  viz.,  Bessemer 
and  non-Bessemer.  This  arises  from  the  fact  that  a  steel  containing  more 
than  0.10  per  cent  phosphorus  is  brittle,  and  hence,  in  the  acid-Bessemer 
steel-making  process,  the  pig  iron  used  must  have  rather  less  than  that 
quantity.  Now  since  about  2  tons  of  iron  ore  make  a  ton  of  pig  iron,  the 
iron  ore  should,  for  safety,  not  contain  more  than  0.045  per  cent  phosphorus. 
At  about  that  limit  the  ore  is  called  non-Bessemer.  In  the  duplex  process 
of  steel  making  this  distinction  has  become  of  little  importance,  since  a 
steel  from  the  Bessemer  converter,  high  in  phosphorus,  can  be  freed  from 
it  in  the  open-hearth  treatment  that  follows. 

Iron  ores,  especially  in  the  United  States,  are  oxides.  They  may  be 
divided  into  hematites,  magnetites,  and  brown  iron  ores.  The  carbonates 
are  there  found  sparingly. 

The  Hematites. — The  iron  in  them  exists  as  F^Oa  (70  per  cent  Fe). 
Of  iron  ores  red  hematite  is  the  most  desirable.  Most  of  the  Lake  Superior 
deposits  are  of  this  variety.  The  ores  from  the  Michigan  and  Wisconsin 
districts  or  ranges  are  called  "  old-range  ores."  Much  hard  or  lumpy  ore 
comes  from  these  ranges  while  from  the  Minnesota  ranges  much  of  the  ore 
is  soft. 

Following  we  give  the  composition  of  some  of  these  ores  as  shipped 
to  the  iron  furnace  in  their  natural  (or  moisture-containing)  condi- 
tion. 

But  Mesabi  ores  will  vary  from  39  to  67  per  cent  in  iron.  The  poor 
ores  are  now  concentrated  to  bring  them  up  to  shipping  grade. 

283 


284 


IRON  ORES  AND  THEIR  SMELTING 


Range. 

H20. 

Fe. 

SiOz 

s. 

P. 

Mn. 

Marquette,  Mich  
Menominee,  Mich 

2.0 
7.5 

56.0 
59.0 

12.0 
4  5 

0.01 
0  01 

0.04 
0  38 

' 

Gogebic,  Wis 

11  0 

54  0 

6  0 

0  03 

0  07 

0  4 

Vermilion,  Minn  

3.5 

64.0 

4.0 

Tr 

0.10 

Mesabi,  Minn.: 
Bessemer  
Non-Bessemer  

60.5 
58.0 

5.2 
6.5 

0.04 
-.09 

0.9 

An  average  analysis  of  Lake  Superior  shipping  ores  in  1919  gave 
H2O,  11.3  per  cent;  Fe,  58.4  per  cent;  SiO2,  7.7  per  cent;  S,  0.06  per 
cent;  P,  0.09  per  cent;  Mn,  0.7  per  cent,  the  bulk  of  the  ore  being  of 
Bessemer  grade. 

Among  the  ores  of  the  United  States  we  should  also  mention  the  Ala- 
bama beds,  and  those  of  Colorado  and  Wyoming  in  the  West. 

There  is  a  red  hematite  at  Sunrise,  Wyo.,  which  carries  Fe,  62  per  cent 
and  beds  at  Orient,  Colo.,  of  easily  reducible  limonite  of  50  per  cent. 

The  Alabama  ores  may  be  divided  into  three  varieties:  First,  the  brown 
ore  or  limonite,  averaging  Fe,  51  per  cent;  P,  0.4  per  cent  and  S,  0.10  per 
cent;  second  the  soft  hematite  of  Fe,  47  per  cent  and  with  as  much  as  17 
per  cent  in  silica;  and  third  the  hard  red  carbonate  ore  of  Fe,  37  per  cent; 
SiO2,  13  per  cent;  P,  0.37  per  cent  C02  and  12.2  per  cent.  They  are  hard, 
heavy  and  hence,  usually  nearly  black  in  color. 

Magnetite  occurs  in  large  deposits  in  Sweden,  and  in  various  parts  of 
the  United  States.  While  some  of  the  beds  are  rich,  many  contain  no 
more  than  40  per  cent  iron  and  carry  so  much  silica  that  the  fluxing  and 
smelting  is  not  profitable.  Much  work  has  been  done  in  the  concentration 
of  these  ores,  both  in  Sweden  and  the  United  States.  In  New  Jersey 
extensive  beds  occur  that  have  been  utilized  by  Edison  for  the  production 
of  a  high-grade  ore  on  a  commercial  scale.  He  has  mined  the  deposit, 
crushed  and  concentrated  it,  and  made  it  into  briquettes  that  contain  as 
little  as  3.3  per  cent  silica  and  0.04  per  cent  phosphorus,  and  as  high  as 
67  per  cent  iron.  The  enterprise,  however,  could  not  continue  at  a  profit 
in  competition  with  foreign  ores  from  Cuba  and  Spain.  Cuban  ores 
include  magnetites  of  Fe,  50  per  cent;  SiO,  10  per  cent;  S,  0.37  orer  cent; 
Cu,  0.10  per  cent.  The  Spanish  spathic  ore  has  H2O  0.91  per  cent;  Fe, 
48  per  cent  and  SiO2,  10  per  cent.  Some  of  the  New  York  beds  near 
Lake  Champlain  have  been  considered  valueless  on  account  of  the 
presence  of  titanium,  it  having  been  asserted  that  this  element  pro- 
duces an  infusible  sticky  slag.  This,  however,  has  been  proved  to  be 
unfounded,  and  it  should  not  prevent  their  use  as  a  source  of  iron.  It 
is  to  be  noted  that  the  famous  Iron  Mountain,  Missouri,  is  a  deposit 
containing  31  per  cent  iron  and  6  per  cent  titanium  oxide,  but  the  deposit 


IRON  ORES  235 

is  not  now  worked.  In  Pennsylvania  the  Cornwall  beds  are  the  most 
important,  and  yield  a  pig-iron  carrying  not  more  than  0.04  per  cent  phos- 
phorus. The  ore  runs  2.5  per  cent  sulphur,  and  about  half  of  this  is 
removed  by  kiln-roasting  before  smelting.  It  contains  also  copper,  which 
will  be  found  in  the  pig  to  the  extent  of  0.5  to  0.75  per  cent.  This  does  not 
matter  in  the  finished  product,  but,  if  the  pig-iron  is  made  into  steel,  the 
copper  causes  "  hot-shortness  "  or  brittleness  when  hot,  thus  causing  im- 
perfections when  rolled  into  shapes.  The  average  of  ore  mined  is  40  to 
42  per  cent  iron  and  20  per  cent  silica. 

At  Fierro,  N.  M.,  is  a  large  deposit  of  hard  magnetite  running  up  to 
61  per  cent  in  iron. 

The  Brown  Ores. — These  are  hydrous  sesquioxides  and  may,  when  pure, 
be  represented  by  the  formula  Fe2O3+3H2O,  equivalent  to  60  to  68  per  cent 
iron.  Limonite,  locally  known  as  bog  ore,  a  brown  hematite,  when  roasted 
has  this  water  of  combination  expelled,  changing  then  to  true  hematite. 
Oolite  is  a  variety  that  exists  in  the  form  of  grains  or  nodules,  and  contains 
silica  and  lime.  When  silicious,  as  in  places  in  Alabama,  the  ore  is  well- 
nigh  worthless,  but  when  limy,  as  in  the  Minette  region  of  Alsace  and  Lor- 
raine, the  ore  is  self- fluxing,  that  is,  the  lime  will  flux  the  silica  that  the 
ore  contains.  A  typical  Minette  ore  carries  Fe,  38  per  cent;  SiC>2,  9.2 
per  cent;  CaO,  12.1  per  cent.  In  the  United  States  these  ores  occur  par- 
allel to  the  Appalachian  range  from  Pennsylvania  into  Alabama,  and 
on  both  sides  of  the  Mississippi,  in  Tennessee  and  Missouri.  Due  to  the 
presence  of  much  phosphorus  these  ores  are  of  non-Bessemer  grade. 
They  will  vary  from  40  to  50  per  cent  in  iron,  5  to  20  per  cent  in  silica, 
0.05  to  0.4  per  cent  in  phosphorus,  and  from  0.3  to  2.0  per  cent  in  mag- 
nesia. 

Carbonate  ore,  siderite  (FeCOa),  as  a  pure  mineral  contains  48.3  per 
cent  iron.  The  varieties  are  spathic,  black  band,  clay-band,  or  clay  iron- 
stone. It  is  often  roasted  to  expel  the  moisture  and  carbon  dioxide  before 
going  to  the  blast-furnaces.  In  England  it  forms  the  well-known  clay 
iron-stone  of  the  Cleveland  district,  but  in  the  United  States,  though  widely 
distributed,  it  is  too  low  in  grade  to  be  used  in  competition  with  the  abun- 
dant rich  ores. 

At  Eisenerz,  Styria,  is  a  celebrated  deposit  of  siderite.  The  ore,  which 
averages  39  per  cent  iron,  is  worked  in  vast  open  cuts.  The  Spanish 
spathic  ore  has  H2O,  9.1  per  cent;  Fe,  48  per  cent  and  SiO2,  10  per  cent. 

ROASTING  IRON  ORES 

In  order  to  expel  moisture,  carbon  dioxide,  and  sulphur,  and  to  render 
the  ore  more  porous,  and  so  more  susceptible  to  reduction  in  the  blast-fur- 
nace, as  well  as  to  decrease  its  weight,  iron  ores,  especially  carbonates,  are 


286 


IRON  ORES  AND  THEIR  SMELTING 


often  roasted,  preferably  ',n  kilns  of  which  Fig.  150  is  an  example.  The 
ore  is  charged  to  the  kiln  in  layers  alternated  with  sufficient  fine  coal,  so 
that  the  roasting  heat  is  maintained  and  the  kiln  is  kept  quite  full.  Peep- 
holes or  openings  at  the  boshed  part  of  the  kiln  are  for  observation  and  for 
loosening  the  charge  by  aid  of  a  bar  as  needed.  Ore  is  unloaded,  as  shown, 
from  cars  into  the  kiln.  The  roasted  material  runs  out  at  the  bottom  as 


VERTICAL  SECTION 


ELEVATION 


FIG.  150. — Gjiers  Calcining  Kiln. 

fast  as  it  is  removed.     Air  is  admitted  to  the  midst  of  the  charge  under  the 
cast-Iron  hood,  at  the  apex  of  the  central  cone. 


THE  AGGLOMERATION  OF  FINE  ORES 

In  modern  blast-furnaces,  due  to  their  high  stacks  and  heavy  blast 
pressure,  it  is  difficult  to  smelt  a  high  percentage  of  dusty  or  finely  gran- 
ulated ore,  such  as  those  of  the  Mesabi  range.  Where  60  or  70  per  cent  of 
such  ores  must  be  used  there  is  a  heavy  flue-dust  loss,  scaffolding  of  the 
furnace,  and  frequent  explosions.  The  fine  ore,  descending  more  quickly 
than  the  rest  of  the  charge,  is  imperfectly  reduced,  causing  disturbances  in 
furnace  operation  and  the  production  of  "  off-iron,"  that  is  of  pig  iron  of 
other  than  the  expected  grade.  Fine  ores,  therefore,  ought  to  be  put  in 
lump  form.  This  may  be  accomplished: 

(1)  By  nodulizing  in  revolving  kilns  or  cylinders  similar  to  Fig.  72  but 
longer.     For  such  ores  as  shrink  much  in  roasting  the  method  is  valuable, 
but  the  product  is  seldom  uniform. 

(2)  By  blast-roasting.     This  is  done  on  a  Dwight-Lloyd  sinter  machine, 


PIG-IRON  SMELTING  287 

as  described  on  page  112.    The  product  is  porous  and  well  sintered  to  blast- 
furnace smelting. 

(3)  By  briquetting.  This  is  done  with  or  without  binding  material, 
and  is  usually  followed  by  the  heating  of  the  briquettes  to  resist  breaking 
and  the  heat  in  the  blast-furnace.  Care  must  be  taken  in  making  them 
that  they  retain  their  porous  character. 

SMELTING  FOR  PIG-IRON 

Outline  of  the  Process. — The  operation  is  conducted  in  a  furnace, 
often  100  ft.  high,  filled  with  a  mixture  of  coke,  iron,  ore,  and  limestone. 
Superheated  air  is  blown  in  at  the  bottom.  The  coke  is  burned  to  main- 
tain a  high  temperature  in  the  furnace  and  to  reduce  the  iron  in  the  ore  to 
the  metallic  form  as  pig  iron.  The  pig  iron  collects  at  the  hearth  or  bottom 


fFiQ.  151. —  Three  Traversing-bridge  Tramways  with  5^-ton  Grab-buckets;    Storage 

capacity,  500,000  tons. 

of  the  furnace,  and  is  removed  from  tune  to  time.  The  gangue,  or  siliciouff 
part  of  the  ore,  is  fluxed  with  limestone,  and  produces  a  worthless  slag,  or 
cinder,  which  is  also  removed  (tapped)  as  it  accumulates  in  the  furnace. 

IRON  BLAST-FURNACE  AND. PLANT 

Blast-furnace  Plant. — Fig.  152  is  a  view  of  an  iron  blast-furnace  plant 
for  the  manufacture  of  pig  iron  from  iron  ores.  In  the  foreground  is  a 
cylindrical  furnace-stack  100  ft.  high,  immediately  in  front  of  which  is  the 
forked  "  down-comer  "  (see  39,  Fig.  155),  a  large  pipe  that  conveys  the 
smoke  from  the  stack  near  the  top  downward  to  the  flue-system  that  car- 
ries it  away.  In  front  of  the  down-comer  is  seen  the  inclined  hoist  for  the 
11  stock  "  or  the  materials  that  are  put  into  the  furnace.  At  the  middle  of 
the  illustration  are  the  four  cylindrical  "  stoves/'  as  high  as  the  furnace, 
used  for  preheating  the  air  blown  into  the  furnace,  while  the  highest  stack 


288 


IRON  ORES  AND  THEIR  SMELTING 


behind  them  draws  away  the  gas  from  the  stoves.  In  front  of  the  four 
stoves  is  the  blast-main,  a  pipe  5  ft.  diameter  by  which  the  air  is  con- 
ducted to  the  furnace.  At  the  left  of  the  stoves  is  the  building  (not  shown), 


FIG.  152.— Blast-furnace  Plant. 

that  contains  the  vertical  blowing  engines  by  which  air,  under  15-lb.  pres-* 
sure,  is  delivered  through  the  stoves  to  the  blast-furnaces.  Fig.  160  rep- 
resents a  vertical  blowing  engine. 


THE  IRON  BLAST-FURNACE  PLANT 


289 


The  general  arrangement  about  the  furnace  is  understood  from  the 
elevation  Figs.  151  and  153  an4  the  plan  Fig.  159.  The  blast-furnace  100 
ft.  high  is  at  the  right.  It  is  served  by  an  inclined  hoist,  one  skip  of  which 
is  hi  position  in  the  pit  ready  for  loading,  the  other  in  discharging  position 
at  the  top.  Within  the  stock-house,  at  the  left,  are  three  standard-gauge 
tracks  on  three  levels.  The  upper  one  at  the  left  is  for  the  hopper-cars 
that  deliver  iron  ore  and  fluxes  to  sloping-bottom  bins  beneath,  shown  in 
section. 

The  second  one  leads  to  similar  bins  (not  shown  in  section),  and  the 
third  to  the  floor  on  the  ground  level.  On  this  level  is  a  charge-car,  elec- 


FIG.  153. — Section  of  Blast  Furnace,  Showing  Filling  Arrangement,  Bins  and  Ore-bridge, 

trically  driven,  with  a  weighing  attachment  that  can  be  brought  to  any 
bin  to  receive  a  weighed  amount  of  stock.  The  load  is  then  transferred 
to  and  discharged  into  the  skip.  In  case  of  accident  to  the  charge-car,  or 
any  trouble  at  bins,  the  furnace  can  be  supplied  by  the  use  of  hand-barrows 
or  buggies,  taking  the  stock  from  the  piles  that  have  been  made  beneath 
the  third  track.  Hoisting  is  done  by  a  hoisting  engine  set  well  out  of  the 
way  at  the  top  of  the  stock-house.  Just  beyond  and  at  the  right  of  the 
furnace  is  the  cast-house,  where  the  molten  iron  is  molded  into  pigs  when  a 
•cast  is  made. 

Iron  Blast-furnace  Plant  at  a  Lake  Port. — We  give  in  Fig.  151  a 
transverse  elevation  of  a  furnace  plant  as  arranged  at  a  lake  port  where 
the  ore  has  been  stored  for  use  in  the  winter  months  and  where  a 


290 


IRON  ORES  AND  THEIR  SMELTING 


traveling  crane  is  used  to  reclaim  it.  The  same  crane  has  also  been 
used  to  unload  from  the  ore  boat  at  its  other  end.  The  ore  is  taken  from 
the  pile  by  means  of  a  10-ton  grab  bucket  which  loads  it  into  cars  standing 
upon  tracks,  these  set  so  they  can  readily  discharge  into  the  feed  pockets 


FIG.  154. — Blast-furnace  with  Automatic  Charging. 

of  the  blast-furnace.  Coke  is  brought  in  directly  from  the  coke  ovens  by 
car.  The  feed  pockets,  as  shown,  are  semicircular  and  both  coke  and 
ore  are  withdrawn  from  them  much  as  shown  in  Fig.  154.  It  will  be  seen 
that  the  charge-car,  which  is  self-weighing,  runs  under  any  of  the  feed 


THE  IRON  BLAST-FURNACE 


291 


pockets,  and  the  materials  weighed  into  it  are  then  taken  to  the  feed  skip 
of  the  furnace. 

The  Iron  Blast-furnace. — Fig.  155  is  a  furnace,  shown  in  part  section, 
and  part  elevation.  It  is  circular  in  cross-section. 

Beginning  at  the  bottom  there  is  a  heavy  foundation  of  concrete  and 
firebrick  upon  which  rests  the  hearth  15  and  columns  4  which  support  the 


FIG.  155. — Blast-furnace,  Detailed  Section. 

upper  brickwork  that  constitutes  the  shaft  of  the  furnace.  The  hearth  or 
crucible  (14J  ft.  diameter  by  9J  ft.  deep),  that  contains  the  molten  iron  and 
slag,  extends  from  the  foundation  to  a  height  slightly  above  the  tuyeres  22. 
The  bottom  and  walls  (see  Fig.  156)  are  of  firebrick,  high  hi  alumina 
like  V  of  page  34,  but  soft  and  porous.  Near  the  bottom  is  the  iron-tap 
27,  through  which  the  molten  pig  iron  is  withdrawn  when  a  quantity  has 


292 


IRON  ORES  AND  THEIR  SMELTING 


accumulated.     At  23  is  the  cinder-notch  or  tap  by  which  the  slag  or  cinder 
is  drawn  off.     The  crucible  is  surrounded  by  a  hearth- jacket  of  steel  plates. 


FTG.  156. — Blast-furnace  Hearth  and  Bosh. 


cooled  on  the  outside  by  sprays  of  water  that  play  against  it,  cooling  and 
protecting  it  and  the  brickwork  lining  from  the  corrosive  action  of  the  mol- 


THE  IRON  BLAST-FURNACE  293 

ten  slag  inside.  Air,  under  a  pressure  of  5  to  14  Ib.  to  the  square  inch, 
enters  through  the  tuyeres  21,  which  have  projecting  nozzles  22,  as  more 
fully  shown  in  Fig.  156.  Care  is  taken  to  withdraw  the  slag  before  it  reaches 
the  level  of  the  tuyeres,  for  it  would  enter  the  openings  and  close  them. 
Of  these  tuyeres  there  are  six.  The  air  is  supplied  through  the  tuyere- 
stocks  33  from  the  bustle-pipe  13,  which  encircles  the  furnace,  and  con- 
nects with  the  blast-main  supplying  air  at  the  temperature  of  a  red  heat 
from  the  stoves.  The  bustle-pipe,  tuyere,  and  tuyere-stock  are  shown  in 
the  section  Fig.  156.  The  bosh,  or  that  part  of  the  furnace  that  widens 
from  14  ft.  6  in.  at  the  hearth  to  22  ft.  in  a  vertical  distance  of  13  ft.,  is  also 
shown.  It  is  in  the  region  of  the  bosh  that  the  formation  of  the  slag 
occurs,  and  the  brickwork  of  the  bosh  is  subject  to  a  slagging  and  scouring 
action  that  tends  to  attack  and  destroy  it.  To  prevent  this,  hollow  water- 
cooled  bosh-cooler  plates  are  laid  in  the  brickwork  of  the  bosh,  making 
rings  around  the  furnace  at  nearly  every  two  feet  vertically.  The  slag 
cuts  into  the  brickwork  nearly  as  far  as  the  inner  ends  of  these  plates,  but 
the  circulation  of  water  within  them  protects  the  adjacent  brickwork 
from  deeper  corrosive  action.  The  bricks  of  the  bosh  are  a  little  harder 
than  in  the  crucible,  but  are  high  hi  alumina  and  porous. 

That  both  the  bustle  pipe  and  the  tuyere  are  lined  with  firebrick 
is  well  indicated. 

The  shaft,  or  main  brickwork  structure  of  the  furnace,  is  carried  by  the 
cast-iron  mantel  5,  resting  upon  the  columns  4.  It  extends  from  the  top 
of  the  bosh  to  the  throat  at  9.  The  upper  part  of  the  furnace  is  closed  by 
a  bell  47  (Fig.  154  shows  a  double  bell) ,  and  the  gas  escapes  at  the  side 
through  the  down-comer  39.  The  in-wall  69,  is  of  a  hard  firebrick  like  IV, 
page  34,  while  the  main  portion  is  of  common  brick  and  is  sheathed  with  a 
shell  46,  of  steel  plates. 

The  tendency  hi  modern  practice  is  to  insert  cooler-plates  extending 
through  the  inwalls  from  the  mantle  upward,  carrying  these  higher  and 
higher  as  the  necessity  for  protection  demands. 

When  in  operation  the  furnace  is  kept  full  to  a  level  just  below  the 
outlet  to  the  down-comer.  This  level  is  known  as  the  stock-line,  and  the 
furnace  at  this  point  is  15  ft.  diameter.  As  the  stock  smelts  and  sinks, 
charges  are  introduced  and  the  stock-line  is  maintained  at  this  level. 

Ordinarily  the  top  of  the  furnace  is  kept  closed  by  the  conical  bell  47, 
which  is  suspended  from  the  ends  of  the  counterweigh  ted  beams  55.  The 
bell  closes  the  bottom  of  a  circular  hopper  48,  into  which  the  charge  in  this 
particular  furnace  is  supplied  by  buggies  brought  up  by  the  elevator  to  the 
upper  or  charge-floor  of  the  furnace  or  "  tunnel-head,"  as  it  is  called.  To 
drop  a  charge  into  the  furnace,  the  outer  end  of  the  lever  is  raised  by  the 
piston-rod  and  piston  of  the  air-cylinder  60.  The  bell  thus  lowered  permits 
the  charge  to  slide  into  the  furnace,  after  which  it  is  immediately  raised 


294  IRON  ORES  AND  THEIR  SMELTING 

to  close  the  opening  and  stop  the  outward  rush  of  smoke  and  gas  that 
mainly  escape  through  the  hood  61.  The  gas,  containing  dust  from  the 
charge,  passes  off  by  the  brick-lined  down-comer  39  to  the  dust-catcher  40 
(where  a  part  of  the  dust  settles),  and  by  the  goose-neck  pipe  41  to  an 
underground  flue  that  leads  to  the  stoves  and  boilers  where  the  gas  is 
burned.  Rising  from  the  down-comer  is  the  bleeder  37,  that  is  used  when 
it  is  desired  to  relieve  the  top  pressure  of  the  gas  rising  from  the  charge. 
It  is  occasionally  used.  Other  openings  are  provided  closed  by  weighted 
doors,  called  explosion  doors,  so  that  in  case  of  a  slip  or  fall  of  material  due 
to  the  giving  way  of  a  scaffold  or  handing  up  in  the  furnaces,  relief  is  given 
to  the  high  pressure  of  gases  suddenly  released.  At  many  furnaces  the 
stock  is  raised  in  hand-barrows  or  charge-buggies  to  the  furnace-top  or 
tunnel-head  51  by  means  of  a  platform  hoist. 

In  Fig.  154  is  shown  the. present  method  of  charging  with  the  inclined 
hoist.  A  double  bell  is  used  to  prevent  the  escape  of  the  gas.  The  charge 
is  dropped  from  the  hoist  into  the  upper  hopper,  where  it  is  retained  until 
the  lower  hopper  is  empty.  The  smaller  upper  bell  is  then  lowered  and 
the  charge  slides  from  the  upper  into  the  lower  hopper,  while  the  upper 
bell  is  closed.  The  hopper  is  then  ready  to  take  another  charge.  The 
charge  in  the  lower  hopper,  when  needed,  is  dropped  into  the  furnace  by 
lowering  the  lower  bell.  It  slides  outwardly  to  the  walls,  forming  a  ring 
or  ridge,  the  stock  in  the  middle  being  a  little  lower  than  at  the  sides. 

However,  the  skip,  dumping  only  in  one  direction,  is  apt  to  deliver  the 
coarser  ore  to  the  opposite  side  of  the  furnace,  so  that  the  blast  comes  up 
more  freely  there,  producing  greater  heats  and  a  "  hot  spot  "  on  that  side. 
To  overcome  this  trouble  distributors  are  used  where  the  charge  in  the 
bell  is  rotated  through  a  varying  arc  which  delivers  it  successively  to  the 
different  segments  of  the  furnace.  Just  beyond  and  below  the  lower  bell 
is  noticed  the  oval  outlet  to  the  down-comer.  The  stock-line  must  be  kept 
below  this. 

The  skips  of  the  hoist  run  in  balance  and  are  charged  as  follows: 
The  charge-car  on  the  ground  level  is  run  to  the  chute  of  an  iron-ore  bin 
to  receive  the  required  weight  of  ore.  It  is  moved  to  the  limestone  bin 
beyond  to  get  the  needed  quantity  of  limestone,  and  then  to  the  skip 
standing  below  in  the  charge-pit,  where  it  is  discharged.  The  skip  is  ^ext 
hoisted  and  dumped,  while  the  empty  one  is  in  position  to  take  the  load  of 
coke.  After  elevating  the  fuel,  a  charge  of  ore  and  flux  goes  next.  These 
charges  alternate  in  the  furnace  and  form  layer  upon  layer. 

The  dimensions  of  a  blast-furnace  are  limited.  The  considerations  are 
as  follows:  The  hearth  should  be  not  more  than  15  ft.  diameter  lest  the 
blast  fall  properly  to  penetrate  to  the  center  and  maintain  intense  com- 
bustion there.  The  slope  or  angle  of  the  bosh- wall  must  be  such  as  to  give 
proper  support  to  the  charge,  which  rests  upon  it,  and  yet  allow  the  solid 


GAS  CLEANING  295 

coke  to  slip  down;  an  angle  of  80°  is  preferred.  The  height  is  limited  to 
the  height  of  the  smelting  zone.  These  conditions  limit  the  diameter  of 
the  bosh  to  22  ft.  From  the  top  of  the  bosh  the  stack  wall  must  decrease 
in  diameter  to  the  throat  to  give  room  for  the  descending  charge  to  swell 
by  reactions  that  occur  in  its  downward  progress.  This  leaves,  at  the 
throat,  a  diameter  suitable  for  the  proper  distribution  of  charge.  Fur- 
naces have  been  built  higher  than  100  ft.,  but  such  height  has  been  found 
to  be  excessive,  especially  for  fine  ores;  and  the  best  practice  calls  for  90  ft. 
or  less. 

GAS  CLEANING 

The  top  gas  coming  away  from  a  blast-furnace,  especially  when  smelting 
fine  ore,  carries  much  dust  caused  by  the  agitation  of  the  blast.  Some  of 
this  is  settled  out  in  the  dust-catcher,  but  the  gas  still  remains  quite  dusty. 
When  the  gas  is  subsequently  burned  at  the  stoves  the  dust  settles  in  the 
checker  work  and  at  the  boilers  it  attaches  itself  to  the  stoves.  If  the 
gas  is  cleaned  it  burns  more  efficiently  and,  moreover,  it  can  then  be 
used  for  driving  a  gas-engine  blower  plant. 

Fig.  157  gives  the  views  of  a  scrubber  plant  for  gas  cleaning  for  stoves 
and  boilers  for  two  furnaces.  Of  the  figure,  TP,  is  a  front  elevation,  X,  a 
side  elevation,  F,  a  separate  elevation  of  the  dust-catcher,  and  Z  a  plan 
view  of  one  of  the  scrubbers,  to  show  the  arrangement  of  the  water  sprays. 
The  gases  from  the  dust-catchers  of  the  two  furnaces  are  united  in  the  7-ft. 
gas  main,  a,  to  go  to  either  of  two  dust- catchers,  6,  each  leading  to  the 
scrubbers,  s  and  s'.  It  should  be  here  noticed  that  one  of  the  scrubbers,  if 
pushed,  will  clean  40,000  cu.  ft.  of  gas  per  minute  while  the  other  may  be 
by-passed  and  used  as  a  spare,  though  for  the  best  work  both  are  used. 
In  each  scrubbing  tower,  14  ft.  diameter  by  74  ft.  high,  are  two  sets  of  ring 
pipes  each  sending  up  jets  of  water,  these  effectively  cleaning  the  ascending 
gas,  wetting  down  the  dust  particles  and  causing  their  fall  to  the  sump 
below.  Here  is  a  goose-neck  siphon  that  permits  the  discharge  of  the 
wetted  particles  or  ore  pulp.  The  cleaned  gas  from  the  top  of  the  towers 
passes  by  the  down-comer,  c,  c,  to  dust  catchers  d,  d,  that  discharge  into 
the  clean-gas  main  E,  for  use  at  the  stoves  and  boilers. 

THE  HOT-BLAST  STOVES 

The  efficient  operation  of  an  iron  blast-furnace  requires  that  the  air 
entering  at  the  tuyeres  be  brought,  generally  to  a  red  heat  (500  to  750°  C.)^_ 
To  do  this  the  furnace  is  equipped  with  three  or  four  (as  in  Fig.  152) 
regenerative  firebrick  stoves  80  ft.  high  and  14  ft.  diameter.  The 
Cowper  stove  (of  Fig.  158),  for  example,  consists  of  a  tight  shell, 
like  a  boiler  shell  of  steel  plates,  lined  with  firebrick,  and  containing  a 


296 


IRON  ORES  AND  THEIR  SMELTING 


checker-work  of  bricks  of  special  shape  laid  in  open  order  so  as  to  have 
numerous  openings  or  passages  from  the  top  to  the  bottom  of  the  stove. 


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The  gas  from  the  furnace,  containing  24  per  centJ^O,  which  in  burning 
supplies  the  heat,  flowing  along  the  underground  flue  from  the  goose- 
neck before  mentioned,  enters  the  stove  g,  while  the  air  is  admitted  at  a, 


THE  HOT-BLAST  STOVE 


297 


the  two  mingling  and  burning  in  the  vertical  circular  flue  /,  and 
heating  the  checker-work  in  their 
passage  to  the  valve  s,  as  shown 
by  the  arrows.  The  gas  thence 
passes  by  an  underground  flue  to 
a  tall  stack,  200  ft.  high,  shown 
behind  the  stoves  in  Fig.  152. 
Gas  having  burned  in  this  way  a 
half  hour,  the  stove  becomes 
heated.  The  valves  g,  a,  and  s 
are  closed,  and  the  cold-air  valve 
at  c  near  the  bottom  of  the  stove 
is  opened,  admitting  air  under 
pressure  from  the  blowing  engine ; 
the  hot-air  valve  h,  at  the  bottom 
of  the  flue  /,  being  at  the  same 
time  opened.  The  air  rises 
through  the  hot  checker-work, 
descends  the  flue  /,  and  passes 
out  at  d  to  the  brick-lined  hot- 
blast  main,  and  to  the  furnace. 

Meanwhile  the  gas  has  been 
turned  into  the  other  stoves,  and 
is  heating  them  in  the  same  way. 
After  the  blast  has  been  received 
in  the  first  stove  a  half  hour  it  is 
turned  into  the  next  heated  stove, 
and  so  on.  The  extensive  sur- 
face of  the  checker-work  serves 
to  absorb  a  large  amount  of  heat 
from  the  burning-gas,  and  to  im- 
part the  heat  subsequently  to  the 
blast-air.  It  will  be  noticed  that 
the  air  enters  at  the  coldest  and 
leaves  at  the  hottest  part  of  the 
stove.  It  flows  in  a  direction 
opposite  to  that  cf  the  burn- 
ing gas,  thus  insuring  the  maxi- 
mum rise  in  temperature.  In 
the  course  of  half  an  hour,  the 
hot  air,  leaving  the  stove,  falls  at  FlG  i58._c0wper  Hot-blast  Stove, 

least  100°  C.  in  temperature. 

Not  all  the  gas  from  the  furnace  is  needed  for  heating  the  stoves,  and  a 


298 


IRON  ORES  AND  THEIR  SMELTING 


portion  is  burned  under  the  boilers  of  the  plant  for  making  steam  for  power. 
The  amount  of  steam  thus  available  is  sufficient  to  run  the  bio  wing- engines, 
hoisting-engines,  hoisting-mechanism,  and  all  machinery  belonging  to  the 
furnace.  At  some  plants  the  surplus  gas,  after  the  stoves  have  been  sup- 
plied, has  been  cleaned  from  dust  and  used  in  gas-engines.  Power  can 
be  gained  in  this  way  and  is  more  available  for  rolling  mill  or  other 
purposes. 

Formerly  the  air  was  heated  in  iron-pipe  stoves,  the  air  circulating 
through  a  nest  of  pipes  inclosed  in  a  furnace  or  brick  heating-chamber. 
These  are  no  longer  used,  but  have  been  supplanted  by  the  regenerative 
stoves  just  described. 

BLAST-FURNACE  AND  ACCESSORIES 

Fig.  159  is  a  plan  of  the  blast-furnaces,  showing  the  course  of  the  gases 
throughout  the  plant.  From  near  the  top  of  the  blast-furnace,  as  well 


Cold  Blast  TTatn 


Blowing  Engine  House 

FIG.  159. — Plan  of  Furnaces  and  Stoves. 

shown  in  Fig.  152,  come  away  the  two  branches  of  the  down-comer  which 
uniting  in  one  enter  the  top  of  the  dust-catcher.  This  settles  out  much  of 
the  flue-dust  in  the  furnace  gases  produced  by  the  heavy  blast  in  the  fur- 
nace. The  rest  is  removed  at  the  gas  washer,  so  that  a  dust-free  gas  may 
be  used  at  the  boilers  and  the  stoves.  A  gas-main,  taken  off  at  the  gas- 
washer,  passes  along  the  front  of  the  boilers  with  branches  to  each,  where  it 
is  burned  for  the  generation  of  the  steam  needed  for  driving  the  blowing 


THE  BLOWING  ENGINE 


299 


engine,  that  taking  some  9  per  cent  of  the  total  heat.  The  other  branch 
from  the  gas-washer  passes  along  in  front  of  the  stoves,  some  14  per  cent 
of  the  total  heat  being  there  utilized.  From  the  blowing-engine  house, 
where  there  are  three  blowing  engines,  comes  away  the  cold-blast  main  to 
the  stoves.  Here  the  air  is  heated  500°  to  750°  C.,  and  goes  by  the  hot- 
blast  main  to  the  bustle  pipe,  thence  through  the  tuyeres,  into  the  blast- 


FIG.  160. — Blowing-engine. 

furnace.     Both  the  hot-blast  main  and  the  bustle  pipe  are  lined  with 
to  prevent  heat  loss  by  radiation  from  the  pipes. 

Blowing  Engines. — Fig.  160  gives  a  view  of  a  vertical  blowing-engine, 
having  an  air-cylinder  at  the  top  7  ft.  in  diameter  and  of  a  5-ft.  stroke. 
The  cylinder  displaces  385  cu.  ft.  air  per  revolution,  or  15,400  cu.  ft. 
free  air  per  minute,  and  delivers  it  at  15-lb.  pressure.  The  air-admission 
and  discharge  valves  are  arranged  to  operate  positively,  and  to  open  and 


300  IRON  ORES  AND  THEIR  SMELTING 

close  at  exactly  the  right  moment.  Enough  of  these  engines  are  installed 
to  supply  the  amount  of  air  required  by  the  furnace. 

OPERATION  OF  THE  BLAST-FURNACE 

The  ore  is  dumped  into  the  furnace  with  45  to  60  per  cent  of  its  weight 
of  coke,  and  the  limestone  needed  to  form  the  predetermined  slag.  The 
furnace  should  be  at  least  65  ft.  high,  and  is  now  built  80  to  100  ft.  high. 
It  is  kept  full  of  stock,  and  the  combustion  of  the  coke  is  supported  by  the 
air  introduced  at  high  pressure  at  the  tuyeres.  As  smelting  progresses,  the 
coke  burns,  the  slag  and  iron  produced  from  the  charge  is  withdrawn,  and 
the  surface  or  stock-line  sinks.  Thus  the  removal  of  molten  products 
below  and  the  addition  of  fresh  stock  above,  cause  the  greatest  production 
of  heat  next  the  tuyeres,  where  the  coke  largely  burns,  the  temperature 
decreasing  toward  the  stock-line.  The  actual  melting  zone,  or  zone  of 
fusion,  extends  upward  through  the  bosh  region.  The  most  intense  com- 
bustion occurs  within  4  ft.  of  the  tuyeres,  where  an  excess  of  air,  driven 
in  under  high  pressure,  burns  the  coke  to  carbon  dioxide.  In  the  reac- 
tion, the  C02  may  be  said  to  dissolve  the  carbon,  it  being  as  follows: 

(1)  C02+C=2CO 

97,000  2  X  29,000  =  -  39,000 

The  reaction  being  endothermic,  lessens  the  temperature  in  this  second 
region.  The  temperature  is  so  high,  however,  that  the  glowing  coke 
reduces  to  iron  any  iron  oxide  that  descends  as  far  down  in  the  furnace  as 
this  region.  The  rising  gas  consists  of  COCO2  and  N.  The  carbon  monox- 
ide, in  the  ascent  reduces  the  iron  oxide  to  iron.  On  reaching  the  lower  part 
of  the  furnace  the  iron,  with  carbon  taken  from  the  CO,  forms  pig  iron. 
The  iron  takes  up  silica,  phosphorus,  and  sulphur  from  the  earthy  constit- 
uents of  the  charge.  The  carbon  amounts  to  3  or  4  per  cent,  and  the  other 
impurities  to  2  per  cent  the  weight  of  the  product.  It  is  the  5  per  cent  of 
the  metalloids  present  in  the  pig  that  makes  it  fusible. 

We  have,  therefore,  in  the  blast-furnace  beginning  from  above,  three 
zones: 

(1)  The  zone  of  preparation,  where  CO2  is  driven  from  the  limestone 
and  moisture  from  the  charge. 

(2)  The  zone  of  reduction,  where  the  CO  of  the  rising  gas  reduces  the 
iron  ore,  first  to  the  ferrous  form,  then  to  iron,  and  where  the  iron  in  a 
spongy  or  open  form  absorbs  carbon  from  the  reducing  gas. 

(3)  The  zone  of  fusion,  where  the  temperature  of  the  furnace  is  high 
and  the  slag  is  formed,  the  iron  at  the  same  time  absorbing  silicon,  phos- 
phorus, and  sulpur. 

In  the  upper  zone  of  the  furnace,  the  carbon  dioxide  of  the  limestone  is 
expelled,  leaving  quicklime  (CaO)  ready  for  fluxing  the  gangue  or  waste 


BLAST-FURNACE  OPERATION  301 

matter  of  the  charge.  The  action  is  endothermic,  lessening  the  heat  of  the 
escaping  gases.  The  quantity  of  limestone  needed  to  form  a  suitable  slag 
is  calculated  in  advance. 

The  gas  varies  in  composition,  but  commonly  contains  61  per  cent 
nitrogen,  from  10  to  17  per  cent  CC>2,  and  22  to  27  per  cent  CO.  The  car- 
bon monoxide  is  the  combustible  constituent  of  gas  that  produces  the  heat 
when  the  gas  is  burned  in  the  stoves  and  boilers. 

In  the  foregoing  description  it  has  been  assumed  that  coke  is  the  fuel 
employed  as  is  true  hi  most  cases.  Anthracite  coal  has  been  used  when 
cheap  enough  to  compete  with  coke,  but  even  then  a  more  satisfactory 
result  is  obtained  when  coke  forms  part  of  the  charge.  Furnaces  in  which  a 
part  of  the  fuel  is  anthracite  are  called  anthracite  furnaces,  but  the  name 
is  somewhat  misleading.  Other  furnaces  use  charcoal  exclusively.  Char- 
coal  is  supposed  to  given  an  iron  of  great  toughness  that  is  particularly 
valuable  for  cast-iron  car-wheels  and  other  castings  requiring  toughness. 
Its  superiority  over  other  kinds  having  the  same  constitution  has  been 
widely  disputed,  but  there  is  testimony  in  favor  of  charcoal-iron. 

Thirty  years  ago  blast-furnace  practice  was  regulated  by  rule-of-thumb 
methods.  They  were  "  born  of  a  bigoted  belief,  on  the  part  of  ignorant 
furnace-men,  that  particular  ores  and  fuel  could  be  worked  in  a  furnace 
only  on  special  lines,  and  that  it  was  impious  to  drive  a  furnace  faster  than 
a  certain  rate  established  by  time-worn  tradition."  In  1879  certain  experi- 
ments made  at  the  Edgar  Thompson  Steel  Works,  Pittsburg,  Pa.,  showed 
that  it  was  possible  to  increase  the  output  of  a  furnace  enormously  by 
increasing  the  air-supply.  It  was  also  found  that  the  amount  of  air,  not 
the  pressure,  determined  the  rapidity.  Under  the  new  system  it  was 
thought  necessary  to  make  a  steep-angle  bosh  (80°)  resembling  that  in 
Fig.  155  more  than  that  in  Fig.  154.  With  the  more  rapid  driving,  reduc- 
tion decreased,  and  the  slag  contained  more  iron.  To  secure  the  reduction, 
the  fuel  had  to  be  kept  high,  using  one  ton  of  coke  per  ton  of  pig  iron  pro- 
duced, and  where  coke  was  expensive  this  was  a  serious  matter.  E.  C. 
Potter  at  the  Illinois  Steel  Works,  South  Chicago,  showed  that  by  reducing 
the  bosh-angle  to  75°  and  using  somewhat  less  blast,  it  was  possible  to 
cut  the  coke  consumption  from  2240  Ib.  to  1800,  or  even  1750,  per  ton  of 
pig  produced.  Furnaces  with  large  hearths  were  then  built,  which  also 
increased  capacity. 

Bio  wing-in. — The  furnace  is  first  dried  several  days  by  a  wood  fire  in 
the  crucible.  The  lower  part,  halfway  up  the  bosh,  is  filled  with  cord-wood. 
Upon  this  is  placed  a  heavy  bed  of  coke  with  limestone  to  flux  the  coke-ash, 
followed  by  successive  layers  of  the  normal  charge  of  coke,  with  gradually 
increasing  amounts  of  ore  and  limestone,  and  decreasing  quantities  of  slag, 
until  the  normal  charge  of  ore  and  flux  is  reached.  The  wood  is  ignited  at 
the  tuyeres,  and  a  weak  blast  of  air  supplied.  The  pressure  is  gradually 


302  IRON  ORES  AND  THEIR  SMELTING 

increased  during  twenty-four  hours,  and  the  furnace  becoming  entirely 
filled  during  this  time  the  regular  pressure  is  reached. 

Regular  Operation. — The  old  way  of  charging  the  furnace  by  hand  is 
as  follows:  Ores,  flux,  and  fuel  (the  "  stock  ")  is  brought  in  buggies  from 
the  stock-house  to  the  scales,  weighed,  hoisted  to  the  top  of  the  furnace,  or 
tunnel-head,  wheeled  by  top-fillers  to  the  bells,  and  dumped  evenly,  the 
fuel  separately,  the  ore  and  "  stone  "  (limestone)  together.  The  furnace 
is  kept  full  to  a  level  just  below  the  outlet  of  the  down-comer.  In  auto- 
matic charging,  as  indicated  in  Fig.  154,  the  stock  is  charged  to  the  skips, 
as  has  been  described,  and  is  hoisted  to  the  furnace-top  and  dumped  into 
the  double  hopper.  No  top-fillers  are  needed,  and  the  bells  are  often 
operated  from  the  ground-level,  so  that  no  attendant  is  needed  at  the  tunnel- 
head. 

IRREGULARITIES  IN  BLAST-FURNACE  OPERATION 

In  the  upper  part  of  the  smelting  zone  semi-fused  material  may  attach 
itself  to  the  walls  of  the  furnace,  and  this  is  the  beginning  of  an  accretion 
which  forms  a  "  scaffold."  This  scaffold  or  hanging  may  gradually  build 
out  until  it  arches  over,  holding  up  the  charge  and  nearly  stopping  the  fur- 
nace. Such  a  mishap  is  more  liable  to  occur  with  fine  ore  such  as  that  of  the 
Mesabi  range.  Sometimes  this  scaffold  may  be  broken  down  by  suddenly 
cutting  off  the  blast  pressure,  and  allowing  the  full  weight  of  the  charge  to 
come  upon  the  obstruction.  If  this  proves  ineffective,  then,  by  cutting  a 
hole  through  the  wall  of  the  furnace,  the  obstruction  may  be  melted  out  by 
the  aid  of  a  blow-pipe  burning  oil  or  gas. 

Sometimes  the  scaffolding  may  give  way  in  part,  causing  slips  by  which 
material  is  suddenly  precipitated  to  the  hearth,  and  there  results  an  upward 
rush  of  gases  resembling  an  explosion.  This  may  do  damage  to  the  charg- 
ing, interrupt  operations  and  throw  stock  out  of  the  top  of  the  furnace. 
Some  furnaces  are  provided  with  explosion-doors  or  valves  which  open 
under  the  sudden  pressure  and  relieve  the  strain. 

So  much  cold  material,  precipitated  toward  the  hearth  by  a  slip,  tends 
also  to  cause  a  "  freezing  "  or  solidification  of  the  slagged  material  near  or 
over  the  tuyeres.  The  solid  layer  may  sometimes  be  broken  away  by 
driving  in  a  steel  bar  to  enable  the  blast  again  to  enter.  This  may  result 
in  the  heating  up  at  this  point,  and  a  final  melting  away  of  the  obstruction. 
Sometimes  it  is  necessary  to  melt  through  the  frozen  material  by  the  aid 
of  a  blow-pipe,  or,  in  extreme  cases,  to  break  through  with  the  aid  of 
explosives. 

Or  the  molten  iron  near  the  metal  notch  may  get  so  cold  as  to  solidify 
so  that  it  becomes  impossible  to  enter.  Then  another  tap-hole  must  be 
made  by  boring  through  into  the  crucible  at  a  higher  level.  When  the 
furnace  is  again  regularly  working  the  heat  gradually  descends  until  the 


DISPOSAL  OF  SLAG 


303 


whole  contents  of  the  hearth  are  melted  out,  when  the  regular  tap-hole 
can  be  again  used. 

Irregularities  in  the  smelting  change  the  character  of  the  iron  made. 
Thus  cold  material  coming  down  to  the  hearth  will  chill  the  smelting  zone, 
and  cause  the  silica  to  be  low  and  the  sulphur  high,  that  is,  will  make  white 
iron  when  gray  or  soft  iron  is  desired. 

DISPOSAL  OF  SLAG  OR  CINDER 

On  account  of  its  low  specific  gravity,  slag  floats  upon  the  iron.  The 
iron  occupies  the  lower  part  of  the  crucible,  and  accumulates  until  it  reaches 


FIG.  161. — Slag  Ladle  and  Locomotive. 

the  tuyeres,  when  it  should  be  drawn  off.  Every  1^  to  two  hours,  the 
plugged  cinder-notch  is  pierced  with  a  pointed  steel  rod,  and  the  cinder 
above  the  level  allowed  to  flow  out.  It  flows  along  a  cast-iron  launder  a 
distance  of  15  to  30  ft.  and  falls  into  a  14-ton  slag-car  standing  on  a  track 
below.  When  loaded  the  car  is  hauled  by  a  locomotive  to  the  dump  that 
may  be  a  mile  away,  and  the  contents  of  the  ladle  is  poured  out  at  the  side 
of  the  track.  The  track  is  gradually  raised  and  moved  outward  toward 
the  edge  of  the  dump  as  it  grows. 

A  large  coke  furnace,  yielding  500  tons  of  pig  daily,  when  smelting  ores 
of  good  grade,  produces  300  tons  of  cinder.  In  smelting  silicious 
ores,  the  quantity  of  slag  may  be  twice  as  great. 


304 


IRON  ORES  AND  THEIR  SMELTING 


DISPOSAL  OF  PIG  IRON 

Every  four  to  six  hours  the  metal  is  tapped  from  the  furnace,  50  tons  at  a 
tune.  The  flow  is  started  by  a  pointed  steel  bar  which  is  driven  through 
the  clay-plugged  iron-notch  or  tap-hole.  A  clay-lined  launder  conducts 
the  flow  to  ladles  similar  to  the  cinder  car.  Distant  from  the  furnace  10 
to  15  ft.  is  a  cross-channel  made  in  the  sand.  The  slag  that  floats  on  the 
stream  of  iron  is  diverted  into  the  cross-channel  with  a  skimmer.  An  iron 
plate  is  placed  across  the  flowing  stream  in  such  a  way  as  to  permit  the 
heavy  iron  to  flow  beneath,  while  the  light  slag  is  diverted.  A  charcoal 
furnace,  having.an  output  of  100  tons  per  day,  produces  so  little  cinder  that 
none  is  tapped  at  the  cinder-notch,  but  flowing  out  with  the  metal  it  is 
skimmed  as  above  described.  It  is  run  outside  the  cast-house  upon  the 
ground,  is  allowed  to  cool,  and  is  then  broken  up  and  carted  away.  The 
iron  contained  in  the  ladles  is  called  "  direct  metal  "  and  may  -be  taken 
to  the  steel  works  and  used  in  molten  form. 

In  practice  elsewhere,  the  iron  is  cast  in  molds  in  the  sand  of  the  floor 


FIG.  162. — Heyl  &  Patterson  Pig-casting  Machine. 

of  the  cast-house.  The  floor,  40  ft.  wide  and  80  to  150  ft.  long,  consists 
of  the  sand  in  which  the  depressions  are  molded  and  connected  by  a  main 
channel  or  runner,  which  receives  the  molten  iron  flowing  from  the  furnace. 
The  molds  fill  successively  and  form  pigs  of  iron  weighing  150  Ib.  each. 

To  reduce  the  cost  of  handling  the  pig  metal,  and  to  give  a  product 
smooth  and  free  from  adhering  sand,  casting  machines  have  been  intro- 
duced in  modern  plants.  Fig.  162  is  a  Heyl  &  Patterson  pig-casting  machine, 
consisting  of  an  endless-chain  conveyor  composed  of  a  series  of  molds, 
each  capable  of  holding  120  Ib.  iron.  The  iron,  brought  from  the 
furnace  in  a  large  ladle  shown  at  d,  Fig.  162,  is  poured  into  the  molds 
as  they  travel  slowly  along.  The  pig-iron  chills  quickly,  and  by  the  time 
it  reaches  the  discharge  end,  it  consists  of  solid  pigs  of  iron  and  drops  into 
the  railroad  car  that  is  placed  in  position  to  receive  it.  The  molds  on  their 
return,  inverted,  take  at  c  a  spray  of  whitewash,  the  water  of  which  quickly 
dries  by  the  heat  of  the  mold.  It  leaves  a  coating  of  lime  inside  that  pre- 
vents the  iron  from  adhering.  Mechanical  casting  has  the  advantage  over 
casting  in  sand-molds  that  it  does  away  with  the  hot  and  severe  work  of 


DRY-AIR  BLAST 


305 


breaking  and  handling  the  pigs.     In  hot  weather  the  work  can  hardly 
be  borne,  and  there  always  is  difficulty  in  getting  or  keeping  the  men. 

Blowing-down. — When  a  furnace  is  to  be  put  out  of  blast,  charging  is 
stopped,  and  a  layer  of  coke  is  added  for  the  last  charge.  With  continued 
blast  the  stock-line  descends  and  the  operation  progresses  as  long  as  iron 
and  cinder  can  be  tapped  out,  the  blast  being  gradually  diminished. 
Finally  the  blast  is  stopped,  and  the  remainder  of  the  contents  is  with- 
drawn through  a  hole  broken  in  the  brickwork  near  the  bottom. 


DRY-AIR  BLAST 

This  aims  to  regulate  the  one  big  variable  in  the  operation  of  the  blast- 
furnace. All  the  other  ma- 
terials, the  ore,  stone,  and  coke 
vary  at  the  most  but  a  few  per 
cent  in  their  composition  from 
time  to  time,  while  the  mois- 
ture in  the  air  may  vary  more 
than  100  per  cent  from  day  to 
day,  and  of  this  air  the  pro- 
duction of  a  ton  of  iron  re- 
quires almost  double  the 
weight  of  the  other  raw  ma- 
terials. 

The  moisture  in  the  atmos- 
phere is  removed  by  refrigera- 
tion previous  to  the  introduc- 
tion of  air  into  the  stoves. 
Any  moisture,  on  entering  the 
furnace  at  the  temperature  at 
the  tuyeres,  is  at  once  dis- 
sociated by  the  intense  heat 
into  the  component  gases  ac- 
cording to  the  reaction. 


(2)     H20  =  H2+ 0  =  58,060, 


I  Gradual  Dissociation 

of  CO 
2 


Air  Blast 
Tuyeres 


Zone  of  Reduction  of 
V  SiO2,  P2O5.MnO.  etc 

/one  of  Detmlphuriiatton 
'  Zone  of  Carburizatioa 

Molten  Slag 

Molten  Iron 


Hearth 


FIG.  163.— Section  of  Blast-furnace  Showing 
Temperatures. 


or  per  pound  of  hydrogen  29,- 
030  pound-calories  per  pound. 
The  amount  of  air  neces- 
sary to  burn  the  fuel  for  smelting   100  Ib.  of  pig  varies  with   the   tem- 
perature of  the  blast,  but  a  fair  average   may  be  taken  at  5300  cu.  ft. 
With  the  amount  of  moisture  contained  under  average  conditions,  as  pre- 


306  IRON  ORES  AND  THEIR  SMELTING 

viously  assumed  for  Pennsylvania  (3.44  grains)  the  total  moisture  will 
then  be: 

5300X3.44 


7000  (grains  per  Ib.) 

1545  calories X 2. 6  =  4020  calories. 

The  above  figure  represents  the  heat  lost  per  hundred  pounds  of  iron 
smelted,  due  to  the  moisture  in  the  atmosphere  or,  expressed  in  coke  con- 
sumption, somewhat  over  50  Ib.  of  coke  per  ton  of  pig  made. 

The  Gay  ley  process,  as  installed  at  several  plants,  has  removed  by 
cooling  to  25°  or  30°  F.  approximately  65  to  70  Ib.  of  water  per  ton  of 
pig  smelted  during  some  of  the  more  humid  months  of  the  year,  a  theo- 
retical saving  in  the  consumption  of  coke  of  some  55  Ib.  per  ton  of  iron, 
but  the  actual  saving  has  proven  far  greater. 

It  will  be  seen  that  the  actual  saving  far  exceeds  any  that  can  be  theo- 
retically accounted  for,  either  by  the  elimination  of  moisture  or  by  the 
rise  of  temperature  in  the  blast. 

There  has  been  much  discussion  regarding  it,  but  probably  the  greatest 
saving  is  in  reality  attributable  to  the  securing  of  uniformly  favorable  oper- 
ating conditions.  Regularity  is  a  prime  essential  for  economical  running 
and  with  these  conditions  uniformly  good  there  is  no  need  for  the  excess 
fuel  necessary  to  provide  for  contingencies. 

It  would  not,  however,  be  the  part  of  widsom  to  assume  that  such 
enormous  saving  could  be  made  at  all  plants  and  under  all  conditions,  but 
it  is  probably  safe  to  say  that  the  average  plant  can  decrease  its  coke  con- 
sumption at  least  12  per  cent  and  increase  the  production  10  per  cent, 
while  in  very  many  cases  it  is  perfectly  feasible  to  raise  these  percentages 
to  14  per  cent  and  12  per  cent  respectively.  This,  of  course,  refers  to 
what  may  be  relied  upon  the  year  through  and  not  merely  for  short  periods 
under  the  stress  of  record  breaking  output, 

CHEMICAL  REACTIONS  OF  THE  BLAST-FURNACE 

A  blast-furnace  may  be  likened  to  an  immense  gas-producer  in  which 
there  is  a  column,  70  ft.  high,  of  alternate  layers  of  coke,  iron  ore,  and  flux. 
The  column  ranges  in  temperature  from  a  heat  that  shows  no  color  at  the 
throat,  to  a  white  heat  at  the  tuyeres. 

The  hot  air  of  the  blast,  entering  at  the  tuyeres,  strikes  the  white-hot 
coke  with  the  immediate  formation  of  CO2  followed  by  an  instantaneous 
reduction  to  CO.  The  air  therefore  need  only  burn  the  fuel  to  CO  as 
indicated  by  the  following  reaction : 

(3)  C + O  =  CO  =  29,000  Calories. 


REACTIONS  OF  THE  BLAST-FURNACE 


307 


Per  pound  of  carbon  burned  2415  pound-calories  are  generated.     Since 
23  per  cent  of  air  is  oxygen,  and  at  the  sea-level  1  Ib.  of  air  equals  12.38 

16     12  30 

cu.  ft.,  we  have  —  X  =72  cu.  ft.  of  air  per  pound  of  carbon,  or  61 

12      0.2o 

cu.  ft.  per  pound  of  coke  of  85  per  cent  carbon. 

Fig.  164  shows  graphically  the  chemical  reactions  under  a  set  of  condi- 
tions assumed,  while  the  temperature  and  places  where  the  reactions  take 
place  are  shown  in  the  section  of  the  furnace  at  the  extreme  left  in  the  dia- 
gram. To  produce  a  ton  of  pig  iron  (2240  Ib.)  there  is  to  be  used  3520  Ib. 


4X1 


356 


/ 


II 


6288      %     • 


FIG.  164. — Chemical  Reactions  of  the  Blast-Furnace. 


of  60  per  cent  iron  ore  containing  3020  Ib.  Fe2O3,  1888  Ib.  coke,  and  1010 
Ib.  limestone. 

At  the  tunnel-head,  the  iron  ore  (Fe2O3)  plunges  into  an  atmosphere  of 
24  per  cent  CO,  16  per  cent  CO2,  and  60  per  cent  of  N  at  a  temperature  of 
260°  C.  Reduction  of  ferric  oxide  to  Fe3O*  by  the  CO  begins  thus: 

(4)         3Fe203     +    CO 
3X199,400       29,000 


2FesO4      H-      C02 
2X270,800       97,000=  4-11,400  Cal. 


The  reaction  is  completed  at  a  temperature  of  450°  C.  when  the  ore  has 
reached  a  depth  of  10  ft.,  shown  at  2  of  the  diagram.  During  this  period 
the  peculiar  reaction  resulting  in  carbon-deposition  begins,  caused  by 
the  reaction  of  the  gas  on  the  ore,  forming  a  deposit  of  soot  or  carbon 
in  the  pores. 


(5) 


2Fe2O3+8CO  =  7CO2+4Fe+C. 


308  IRON  ORES  AND  THEIR  SMELTING 

Continuing  the  descent  the  ore  undergoes  further  reduction.  At  a 
depth  of  19  ft.  and  a  temperature  of  600°  C.,  the  FesCU  formed,  as  shown 
above,  has  become  further  reduced  to  FeO,  as  indicated  in  column  4,  the 
reaction  being  as  follows : 

(6)  Fe304  +  CO      =  3FeO     +     CO2. 

265,800     29,000     3X66,400     97,000  =+ 1400  Cal. 

The  FeO  thus  formed,  impregnated  with  carbon  (see  column  7)  descends 
with  little  change,  until  at  a  depth  of  26  ft.  and  at  a  temperature  of  700°  C., 
the  CO  of  the  gas  reacts  upon  it,  and  spongy  iron  begins  to  form.  The 
reaction  is  complete  at  800°  C.,  and  at  a  depth  of  32  ft.,  and  is  as  follows: 

(7)  FeO  +   CO     =     Fe+CO2. 

66,400     29,000  97,000  =  + 1600  Cal. 

In  the  passage  downward  the  limestone  gradually  loses  its  CO2,  and 
at  this  point  the  expulsion  is  complete,  as  indicated  at  column  8.  The 
quicklime,  column  9,  thus  formed,  unites  at  the  zone  of  fusion  and  fluxes 
the  silica  of  the  charge.  From  the  depth  of  19  ft.  to  the  depth  at  which 
all  carbon  dioxide  is  expelled,  that  is  between  the  temperatures  550°  and 
880°  C.,  the  CO2  reacts  upon  coke,  dissolving  it  according  to  the  following 
reaction : 

(8)  CO2+C  =  2CO 

97,000       2  X  29,000  =  -  39,000  Cal. 

Thus  heat  is  absorbed  from  the  gas,  and  some  coke  is  consumed.  The 
coke,  however,  as  can  be  seen  from  column  6,  remains  but  little  changed 
until  it  reaches  the  region  of  the  tuyeres. 

Below  the  32-ft.  level  at  800°  C.,  reactions  practically  cease,  the  chief 
action  now  being  a  reduction  of  a  small  amount  of  FeO,  left  undecomposed 
by  the  CO.  This  is  gradually  reduced  (see  column  4)  by  the  glowing  coke 
as  follows: 

(9)  FeO+C  =  Fe  +  CO. 

66,400  29,000  =  -37,400 

Silicon  having  less  affinity  for  oxygen  than  carbon  at  a  high  tempera- 
ture is  formed  from  the  reduction  of  the  silica,  and  as  a  metalloid  enters 
the  pig  iron  after  the  following  reaction : 

(10)  Si02+2C  =  2CO+Si. 


HEAT  BALANCE  OF  THE  BLAST-FURNACE  309 

Below  the  32-ft.  level,  the  temperature  rises  gradually  and  uniformly 
until  the  intense  combustion  at  the  tuyeres  produce  1500°  C.  as  a 
maximum. 

Of  the  air  entering  the  furnace,  77  per  cent  is  nitrogen,  and  of  the 
escaping  gas  60  per  cent,  thus  showing  nitrogen  to  be  by  far  the  largest 
constituent  present.  As  is  shown  in  column  12,  nearly  three  tons  of  nitro- 
gen pass  through  the  furnace  for  each  ton  of  pig  iron  produced.  At  the 
high  temperature  of  the  lower  part  of  the  furnace,  potassium  cyanide  is 
formed,  the  potash  of  the  coke-ash  uniting  with  carbon  and  nitrogen  to  form 
the  salt.  It  decomposes  before  the  top  of  the  charge  is  reached. 

Referring  to  columns  10  and  11,  we  note  that  the  CO2  formed  so  freely 
at  the  tuyeres,  is  at  once  (column  10)  changed  to  CO.  The  carbon  monox- 
ide rises  unchanged  until  it  reaches  the  32-ft.  level,  when  it  begins  to  act 
on  the  iron  oxides  with  the  formation  of  CO2.  The  carbon  dioxide  from 
this  source  united  with  that  from  the  limestone  is  the  total  of  the  escaping 
CO2  gas. 

Sulphur  occurring  as  FeS  in  the  coke  and  as  pyrite  in  the  ore,  is  speedily 
driven  off  by  the  heat  of  the  furnace,  giving  FeS.  The  sulphur  of  the  FeS 
is  taken  up  by  quicklime,  and  enters  the  slag  as  calcium  sulphide  according 
to  the  following  reaction: 

(11)  FeS+CaO+C  =  CaS+Fe+CO 

Thus  it  is  separated  from  the  iron,  upon  which  it  would  have  an  injurious 
effect. 

THE  HEAT  BALANCE  OF  THE  BLAST-FURNACE 

The  heat  yielded  by  the  fuel  and  blast  on  the  one  hand,  and  that 
absorbed  in  the  various  reactions,  taken  by  the  blast,  and  lost  by  radiation, 
may  be  stated  for  a  particular  case  as  follows: 

To  produce  a  ton  (2240  Ib.)  of  pig  iron  of  the  composition,  carbon  42 
per  cent,  silicon  1.35  per  cent,  manganese  0.64  per  cent  and  iron  93.6  per 
cent,  there  was  needed  4093  Ib.  of  ore,  795  Ib.  of  limestone,  and  1682 
Ib.  of  coke,  and  besides  the  2240  Ib.  of  pig  iron,  there  was  yielded  1010  Ib. 
of  slag  and  89  Ib.  of  flue-dust.  There  was  blown  into  the  furnace  6673 
Ib.  of  air  (say  80,000  cu.  ft.)  and  there  came  away  9309  Ib.  of  top-gas  of 
the  composition  by  weight  CO2,  22.3  per  cent;  CO,  22.4  per  cent;  EfeOCHU, 
0.1  per  cent  and  nitrogen  54.9  per  cent.  The  coke  contained  89  per  cent 
of  fixed  carbon  and  9.4  per  cent  ash,  while  the  composition  of  the  ore 
(including  a  little  scrap  returned)  was  51.2  per  cent  iron  and  0.74  per  cent 
manganese.  The  specific  heat  of  air-blast  was  0.248,  the  air  carrying  5.5 
grains  of  moisture  per  cubic  foot,  and  the  average  blast  temperature  was 
672°  C.  The  heat  balance  sheet  for  1  ton  of  pig  iron  is: 


310  IRON  ORES  AND  THEIR  SMELTING 

Generated  by 

Cal. 

Combustion  of  carbon  to  CO 2,220,000 

Combustion  of  carbon  to  CO2 3,752,000 

Heat  content  of  blast  air 1,147,000 

Heat  content  of  moisture  in  air 20,000 


7,142,000 

Consumed  by 
Cal. 

Reduction  of  Fe2O3 3,407,000 

Reduction  of  Fe3C>4 272,000 

3,679,000 

Reduction  of  MnO 27,000 

Reduction  of  SiO2 232,000 

-3,938,000  Cal.  or    55.3% 

Calcination  of  carbonates 400,000  Cal.  or      5.5% 

Dissociation  of  moisture  in  the  blast 220,000  Cal.  or      3.1% 

Carried  off  with  the  iron 635,000  Cal.  or      8.9% 

Carried  off  with  the  slag 550,000  Cal.  or      7.1% 

Carried  off  with  the  dry  top-gas 417,000  Cal.  or      5.9% 

Carried  off  with  the  moisture  in  the  top-gas 388,000  Cal.  or      5.4% 

Radiation,  cooling-water  and  unaccounted  for 631,000  Cal.  or      8.8% 


Total  calories 7,179,000  Cal.  or  100.0% 

It  should  also  be  noted  that  the  heat  carried  off  by  1  Ib.  of  iron  is  261  cal- 
ories, and  that  by  1  Ib.  of  slag  483  calories.  By  a  careful  study  of  the 
above  figures  one  may  arrive  at  a  just  estimate  of  the  value  of  the  various 
furnace  operations,  and  so  can  compare  them  with  the  performance  of 
other  furnaces. 

BURDENING  THE  BLAST  FURNACE 

This  involves  the  calculation  of  the  proper  proportion  of  ore  and  flux 
needed  for  the  production  in  the  furnace  of  a  slag  of  suitable  composition. 
In  order  to  accomplish  this  we  must  have  an  analysis  of  the  materials  of 
the  charge  and  of  the  fuel.  The  furnace  must  work  freely  and  regularly, 
and  must  produce  the  kind  of  iron  desired  as  more  particularly  shewn  on 
page  315.  This  is  accomplished  by  so  burdening  the  furnace  as  to  produce 
a  slag  of  the  proper  composition,  and  by  properly  regulating  the  quantity 
of  fuel  and  the  temperature  of  the  blast. 

The  Slag. — This  results  from  the  melting  together  of  the  non- volatile 
solid  constituents  of  the  charge,  that  is,  the  silica  and  bases  of  the  ore  and 
fluxes,  since  slags  are  essentially  silicates  of  these  bases. 

Below  we  give  the  composition  of  typical  slags  that  have  proved  alto- 
gether satisfactory  in  practice. 


CHARGE  CALCULATION  FOR  THE  BLAST-FURNACE 


311 


SLAG. 

IRON. 

SiO* 

Al,0, 

CaO 

MgO 

CaO  and 
MgO 

Si 

S 

Averages  for  Hot  Furnaces. 


Cuban  ore  

33.2 

13.7 

40.7 

11.1 

51.8 

3.8 

tr. 

Spanish  ore  

34.8 

11.7 

41.3 

9.8 

51.1 

2.5 

0.02 

Spanish  ore  

31.8 

12.0 

45.6 

9.0 

54.6 

1.3 

0.02 

Lake  ore  

35.5 

12.0 

40.5 

8.9 

49.4 

1.8 

0  03 

Averages  for  Cool  Furnaces 


Cuban  ore 

32  2 

10  3 

45  6 

9  8 

55  4 

0  9 

0  07 

Spanish  ore  

30.7 

11.3 

47.4 

8.4 

55.8 

0.4 

0  03 

Lake  ore  

34.7 

11.3 

40.1 

10.9 

51.0 

0.8 

0.06 

Lake  ore  

35.0 

11  4 

39.1 

11.3 

50.4 

0.6 

0.10 

In  these  slags  the  ratio  of  SiO2  to  CaO + MgO  will  average  as  33  to  53 
or  as  1  to  1.6.  Alumina  is  not  regarded  as  an  acid  or  a  base,  but  as  a  neu- 
tral constituent  dissolving  in  the  slag.  As  seen  in  the  table  a  hot-running 
furnace  reduces  more  silicon  to  enter  the  pig,  producing  gray  or  soft  or 
foundry  pig;  when  run  cool  more  like  the  gray  forge,  the  mottled  or  white 
iron ;  also  the  retained  sulphur  in  the  pig  is  higher. 

The  weight  of  the  fuel  of  charge  for  a  good-sized  furnace  should  be  such 
as  will  fill  a  skip  of  5  long  tons  capacity,  the  iron  ore  and  "  stone  "  (lime- 
stone) being  separately  hoisted  and  put  in  to  the  furnace.  These  two  skip 
loads  are  called  a  "  round."  About  a  ton  of  coke  is  needed  to  produce  a 
ton  of  pig,  though  for  a  cool  furnace  as  little  as  1600  to  1800  Ib.  of  coke 
has  been  used.  Two  tons  of  iron  ore  of  50  per  cent  Fe  should  yield  one  ton 
of  pig.  We  may  decide,  that,  as  experience  suggests,  we  need  one-fourth 
of  its  weight  of  limestone.  On  this  basis  we  will  prepare  the  charge  sheet 
as  shown  on  page  312. 

The  amount  of  the  items  are  written  in,  with  their  percentage  compo- 
sition. The  corresponding  weights  of  the  elements  are  calculated  through 
and  their  totals  obtained.  The  11,800  Ib.  of  iron  is  to  have  1.5  per  cent  its 
weight  of  silicon,  or  175  Ib.  Now  1.5  per  cent  Si  corresponds  to  ff  of  silica, 
or  374  Ib.,  and  this  subtracted  from  the  total,  leaves  1589  Ib.  silica  for  the 
slag.  This  multiplied  by  1.6,  as  already  explained,  gives  us  2534  Ib.  Sub- 
tracted from  the  total  it  shows  an  excess  of  672  Ib.,  or  of  limestone  twice 
that,  say  1300  Ib.,  so  only  4200  Ib.  is  needed.  Erase  where  needed,  put 
in  the  new  figure  of  4200  Ib.,  recalculate  and  obtain  new  results,  which 
should  be  nearly  correct.  If  not,  a  second  correction  can  be  carried  out. 


312 


IRON  ORES  AND  THEIR  SMELTING 
CHARGE   SHEET 


Weight. 

Si02. 

Fe. 

CaO-MgO. 

P. 

Per 

Cent. 

Pounds. 

Per 
Cent. 

Pounds. 

Per 

Cent. 

Pounds. 

Per 

Cent. 

Pounds. 

Gogebic  iron  ore..  .  . 
Limestone  
Coke 

22,000 
5,500 
11,000 

6.3 

1.5 
5.4 

1,286 
83 
594 

1,963 
374 

53.5 

11,770 

0.6 

52.7 
1.6 

132 

2,898 
176 

3,206 
2,534 

0.01 
0.04 

=  need 
=  exces 

2.2 
2.2 

4.4 
ed 

s 

11,770 

Pig 

Fe    94.5  
Si        1  5 

For  pig   = 
For  slag  = 

1,589 

672 

C       3.5  
99.5% 

Where  manganese  exists  in  the  ore,  about  one-third  of  it  enters  the 
slag  as  MnO,  making  it  more  fluid.  The  rest  accompanies  the  pig  iron. 
A  little  iron  may,  as  FeO,  enter  the  slag,  but  this  loss  to  the  pig  iron  is  more 
than  made  up  by  the  additions  of  carbon  and  silicon  that  it  receives. 


GENERAL  ARRANGEMENT  OF  THE  BLAST-FURNACE  PLANT 

Fig.  165  shows  the  general  arrangement  of  a  two-furnace  plant.  All 
parts  of  the  plant  are  reached  by  railroad  tracks  on  short  easy  curves. 
The  ore  is  brought  in  by  ore-cargo  steamers  and  stored  for  the  winter 
period  in  an  extensive  ore  storage  yard.  From  the  yard  it  is  reclaimed  by  a 
traveling  crane  and  stored  in  ore  pockets  adjoining  the  furnaces  and  next 
to  the  coke  pockets.  The  coke  is  brought  in  by  an  overhead  track  and 
unloaded  to  the  pockets.  The  blast-furnaces,  marked  respectively  A  and 
B,  have  each  their  cast-house  where  the  iron  can  be  cast  in  sand-beds.  At 
the  end  of  the  building  is  the  pig-breaker  where  the  pig  is  broken  in  order 
to  determine  by  fracture  its  grade.  Between  the  furnaces  are  seen  the 
stoves,  four  for  each  furnace.  The  large  building,  between  the  cast  houses, 
A  and  B,  may  be  called  the  power  house,  and  contains  the  engines,  with 
their  blowers  and  the  boilers  heated  by  furnace-gas,  also  the  machirte  shop. 
To  the  right  of  the -yard  is  the  building  for  the  pig-casting  machine.  The 
molten  pig  iron  from  the  furnaces  is  brought  by  ladle  car  to  the  "  ladle 
house  "  adjoining  the  pig-casting  house  and  there  poured  into  the  molds  of 
the  machine.  Alongside  each  cast-house  is  the  hot  cinder  track  where  the 
cinder-car  is  run  in  to  receive  the  molten  slag  as  it  is  tapped  from  the 
furnace,  and  thence  taken  to  some  distant  dump. 


IRON  BLAST-FURNACE  PLANT 


313 


314  IRON  ORES  AND  THEIR  SMELTING 


^  PIG  IRON 

The  iron  produced  in  the  blast-furnace  is  not  pure,  but  contains  3J  to 
4  per  cent  carbon  and  1J  to  3  per  cent  silicon.  Some  of  the  carbon  is 
combined  chemically,  some  separated  as  graphite.  If  a  large  proportion 
is  combined,  the  metal  is  hard  and  the  fracture  of  the  iron  looks  white. 
If  a  large  proportion  is  free,  the  fracture  is  gray  or  black  with  scales  of 
graphite,  and  the  iron  is  soft  and  tough. 

Under  "  Chemical  Reactions  of  the  Iron  Blast-furnace,"  Equation  (3) 
indicates  the  formation  of  carbon,  and  Equation  (8)  the  reduction  of  silica 
to  silicon,  both  elements  entering  the  pig  iron.  A  small  amount  of  sulphur, 
seldom  less  than  0.2  and  often  0.25  per  cent  or  more,  is  present.  As  the 
amount  increases  above  0.1  per  cent  the  iron  becomes  harder  and  more 
brittle. 

The  percentage  of  silicon  and  sulphur  in  the  iron  depends  in  large  meas- 
ure upon  furnace-conditions;  hence  it  can  be  controlled;  but  all  the  phos- 
phorus present  enters  the  pig  iron.  In  pig  iron  for  steel  manufacture  by 
the  usual,  or  acid  Bessemer  process,  the  phosphorus  in  the  pig  must  not 
exceed  0.10  per  cent.  Therefore,  in  the  ore,  it  must  not  be  higher  than 
0.05  or  0.06  per  cent.  In  the  acid  Bessemer  process  the  phosphorus  is  not 
eliminated,  and  it  tends  to  make  steel  red-short  (brittle  when  hot)  a 
quality  that  interferes  with  the  subsequent  rolling.  Phosphorus,  on  the 
other  hand,  imparts  the  quality  of  fluidity  to  cast-iron.  Iron  that  con- 
tains 3  per  cent  P  is  in  demand  where  intricate  castings  are  to  be  made,  and 
can  be  used  where  brittleness  is  of  minor  importance. 

Cast  iron  as  compared  with  steel  and  wrought  iron  has  the  following 
characteristics : 

(1)  It  is  brittle  because  of  the  presence  of  the  metalloids,  carbon  and 
silicon. 

(2)  Because  of  the  presence  of  the  metalloids  it  is  fusible;  and  it  derives 
thus  the  most  valuable  property.     It  runs  freely  from  the  blast-furnace  and 
can  be  cast  in  intricate  molds  to  form  castings  of  any  kind.     Wrought  iron 
at  the  same  temperature  would  be  pasty  and  would  not  run.     Steel,  which 
is  intermediate  between  wrought  iron  and  cast  iron  in  the  contained  carbon, 
can  be  made  into  castings,  however,  but  not  readily  like  cast  iron.     The 
making  of  steel  castings  is  becoming  more  common. 

(3)  It  cannot  be  forged  either  hot  or  cold. 

CLASSIFICATION  OF  PIG  IRON 

Pig  iron  for  further  treatment  or  use  may  be  thus  distinguished : 
Mill  Iron. — For  puddling,  a  pig  low  in  silicon  is  needed,  but  otherwise  of 
a  different  quality. 


CLASSIFICATIONS  OF  PIG  IRON 


315 


Bessemer  Pig. — By  this  is  meant  an  iron  containing  less  than  0.10 
per  cent  phosphorus  and  less  than  0.05  per  cent  sulphur. 

Basic  Iron. — This  should  be  low  in  silicon,  that  has  been  cast  in  an 
endless-chain  casting-machine,  thus  being  free  from  sand.  Silica  attacks 
the  lining  of  a  basic-lined  open-hearth. 

Malleable-iron  Pig. — Used  in  making  malleable  iron  castings.  It  is 
non-Bessemer,  low  in  silicon  and  graphitic  carbon. 

Charcoal  Iron. — This  is  made  in  charcoal  furnaces,  and  is  used  for 
special  purposes  in  the  foundry — as,  for  example,  in  making  car  wheels. 

Foundry  Pig. — Is  used  for  making  castings  for  all  purposes.  The  iron 
should  readily  fill  the  mold  and  not  shrink  much  when  cast.  Otherwise  a 
grade  of  iron  is  used  to  suit  the  purposes  to  which  the  casting  is  to  be 
put. 

Grading  Pig-iron  by  Fracture. — The  pigs  are  broken  in  two,  either 
over  a  wedge-shaped  block  or  in  a  machine,  and  the  fracture  is  observed. 
Foundry  No.  1  is  dark  gray  in  color,  the  grain  large  and  even;  foundry 
No.  2  has  a  small  uneven  grain  and  is  lighter  in  color;  foundry  No.  3  is 
close-grained  and  light  in  color  but  has  less  than  3  per  cent  silicon,  in  fact 
a  white  iron. 

Grading  by  Analysis. — This  is  a  more  reliable  method  than  by  fracture. 
The  character  of  the  grades  of  Alabama  pig  iron  is  indicated  by  the  table 
below: 


ALABAMA  PIG  IRON 


Name  of  Iron. 

Graphite  Carbon. 

Combined  Carbon. 

Silicon. 

Silver  gray 

3    13 

0   02 

5  5 

No.  1  soft  

3  48 

0  03 

3  5 

No.  2  soft  

5  53 

0  03 

3  5  to  4  0 

No   1  foundry 

3  49 

0  07 

2  8  to  3  5 

No.  2  foundry  

3  55 

0  07 

2  2  to  2  6 

No.  3  foundry  
Gray  forge  ... 

3.48 
3  00 

0.10 
0  57 

2.0  to2.4 
1  3  to  1  7 

Mottled 

2  11 

1  22 

1  1  to  1  6 

White 

0  10 

2  92 

0  7  to  1  2 

This  table  shows  the  increase  of  combined  carbon,  and  the  decrease 
of  silicon,  as  the  grade  approaches  white  iron. 

The  first  grades  are  more  difficult  to  make,  and  command  a  higher 
price. 

Pittsburg  Pig  Iron. — For  the  Pittsburg  district  we  give  a  similar 
table  but  with  the  silicon,  on  the  whole,  lower  and  disregarding  the 
carbon. 


316 


IRON  ORES  AND  THEIR  SMELTING 


Name  of  Iron. 

Silicon. 

Sulphur. 

Phosphorus. 

Manganese. 

Gray  forge 

0  75  and  over 

Over  0  .  05 

0  .  40  to  0  .  60 

0  .  50  to  0  .  80 

Basic  (chill  cast)  .... 
Strong    foundry    and 
car-wheel  
Bessemer  
Low  phosphorus  

Less  than  1  .  00 

0.75  to  1.50 
1.00  to  2.  00 
1.00  to  2.00 

0  .  05  and  less 

0  .  05  and  less 
0  .  05  and  less 
0  .  035  and  less 

0.03  and  less 
0  .  10  and  less 
0  .  035  and  less 

No.  2  foundry  

1.75  to  2.  25 

0  .  05  and  less 

over  1  .  00 

0.35  to  0.70 

INFLUENCE  OF  ITS  CONTAINED  ELEMENTS  ON  THE  CHARACTER  OF  THE 

PIG-IRON 

Carbon. — This  occurs  graphitic  and  combined,  in  ordinary  pig  up  to 
4.5  per  cent,  and  in  high  manganese  and  chrome  iron  to  as  much  as  7 
per  cent. .  When  molten  the  carbon  is  regarded  as  being  combined,  but 
in  cooling  more  or  less  separates  as  graphitic  carbon.  When  much  of 
the  latter  separates  the  fracture  is  darker  in  color  and  softer  than  where 
the  carbon  remains  combined;  it  is  well  suited  to  machining,  though  not 
so  strong  as  when  the  carbon  is  combined.  In  smelting,  to  obtain  a  pig 
that  will  be  high  in  graphite,  the  temperature  of  the  blast  and  the  propor- 
tion of  fuel  should  be  high,  so  as  to  secure  a  good  reduction,  and  that 
much  carbon  shall  be  taken  up  from  the  fuel.  This  also  ensures  reduction 
of  much  silicon  to  enter  the  pig,  and  cooling  compels  carbon  to  take  the 
graphitic  form. 

Silicon. — This  is  reduced  from  the  silica  of  the  charge  at  the  hearth  of 
the  blast-furnace.  It  then  dissolves  in  the  forming  iron.  For  a  high- 
silicon  iron  a  high  temperature  is  here  needed  and  this  is  accomplished 
by  a  light  burden  and  a  good  hot  blast.  Iron,  containing  as  much  as  20 
per  cent  silicon  has  been  made  in  the  blast-furnace,  and  when  the  pig  con- 
tains more  than  6  per  cent  it  is  called  ferro-silicon,  a  product  much  used  in 
steel  making. 

Phosphorus. — All  of  this  element  present  in  the  ore  is  readily  reduced 
in  the  blast-furnace  to  a  phosphide  which  combines  with  the  iron.  It 
makes  the  iron  more  fluid  so  that  it  better  fills  the  mold  but  the  casting  is 
more  brittle. 

Manganese. — This  is  reduced  like  the  iron.  It  aids  the  pig  iniholding 
the  carbon  in  combined  form.  Manganese  increases  the  strength  and 
fluidity  of  the  pig,  and  makes  it  harder  and  less  fusible.  Likewise  it  tends 
to  remove  oxygen  and  sulphur  from  the  iron,  and  to  counteract  the  detri- 
mental effect  of  other  impurities.  To  make  ferro-manganese  alloys,  much 
used  in  steel-making,  a  separate  blast-furnace  is  operated,  using  a  very  hot 
blast  and  a  light  burden  (a  high  fuel),  since  manganese  is  difficult  to  reduce. 
An  alloy  containing  10  to  25  per  cent  manganese  is  called  spiegeleisen 


CHARACTERISTICS  OP  PIG  IRON  317 

(mirror-iron)  because  of  its  shining  crystalline  appearance;  and  when  con- 
taining 25  to  95  per  cent  manganese  it  is  known  as  ferro-manganese. 
Sulphur. — This  detrimental  element  occurs  dissolved  hi  the  pig  metal 
as  FeS.  It  makes  it  hard  and  brittle  and  tends  to  keep  the  carbon  in 
combined  form.  A  high  percentage  of  sulphur  makes  porous  castings, 
but  the  iron  is  more  fluid  when  cast.  In  the  blast-furnace,  using  a  high 
lime  slag,  this  tends  to  take  it  away  from  the  iron. 


CHAPTER  XXVI 
WROUGHT  IRON  AND  STEEL 

Cast  iron,  because  of  the  large  proportion  of  contained  metalloids, 
which  indeed  makes  it  fusible,  and  easily  cast,  is  too  weak  and  brittle  for 
many  structural  purposes. 

Therefore  three-fourths  or  more  of  the  pig  iron  in  the  United  States, 
together  with  much  steel  scrap,  is  made  into  steel  because  of  its  superiority 
as  an  engineering  material.  About  3  per  cent  of  our  pig  iron  is  made  into 
wrought  iron,  a  product  superior  to  steel  for  certain  purposes,  because  of 
its  welding  quality  and  ductibility  as  compared  with  ordinary  Bessemer  or 
open-hearth  steel.  For  most  engineering  purposes  steel  is,  however, 
superior  to  wrought  iron  and  as  cheap. 

THE  MANUFACTURE   OF  WROUGHT  IRON  BY  THE  PUDDLING  PROCESS 

Almost  all  wrought  iron  manufactured  in  the  United  States,  about  1| 
million  tons  per  annum,  is  made  from  pig  iron  by  the  puddling  process, 
invented  in  England  by  Henry  Cort  about  1780,  and  greatly  improved  by 
Joseph  Hall,  fifty  years  later.  The  grade  of  pig  used  is  either  gray  forge 
or  white — see  page  315.  Sulphur  should  not  exceed  0.10  per  cent  and  phos- 
phorus should  preferably  be  less  than  1.0  per  cent.  Pig  containing  as 
much  as  S,  0.35  per  cent  and  P,  2.5  to  3.0  per  cent  is  sometimes  used, 
since  a  high  phosphorus  in  the  resultant  wrought  iron  is  not  so  objection- 
able as  it  would  be  in  steel.  The  slag,  mechanically  mingled  with  wrought 
iron,  hinders  it  from  becoming  brittle  under  shock,  the  difficulty  produced 
by  phosphorus  in  steel. 

The  Furnace. — Fig.  166  is  a  longitudinal  section  of  a  puddling  furnace 
of  about  1500  Ib.  per  charge  capacity.  It  is  a  reverberatory  furnace  hav- 
ing a  hearth  of  7  ft.  by  7  ft.  in  size,  the  grate  being  2  ft.  10  by  4  ft.  fcy  9  in. 
size,  and  relatively  large  for  so  small  a  hearth,  in  order  to  obtain  a  high 
furnace  temperature.  The  hearth  lining  of  mill  cinder  and  iron  ore  suffers 
wear  and  is  repaired  between  the  heats. 

Puddling. — The  pig  is  charged  by  hand  into  the  furnace,  and  is  rapidly 
melted  down  in  thirty  to  thirty-five  minutes.  Iron  ore  or  mill  scale 
(FeaCU)  is  now  added,  this  taking  seven  to  ten  minutes,  and  the  charge  is 
thoroughly  mixed  and  cooled  to  a  point  where  the  slag  will  begin  to  oxidize 

318 


PUDDLING  FURNACE  REACTIONS 


319 


impurities,  especially  phosphorus  and  sulphur.  As  this  takes  place,  a 
light  flame  begins  to  break  through  the  slag-covering,  due  to  the  carbon 
of  the  pig  iron  reacting  on  the  oxides  of  the  bath  thus, 


(12) 


Fe2O3  +3C  =  SCO +2Fe. 

20' 


imitmmm?  m 


:::  M  i  1 1  ;:C  r  u  d  e  r.  ;\  lul :  Iron;  Ore 
•:•':":  .":"•.*: 


Fire  Hole 


Cast Jroii  Plate  '       "2' 

T.~-  -——.r:  -.——.- — — .-•—.— .—  ~  ~  ~  -  -•—  — . — ——.-.——— -~~ 

Ait  Chamber  Air  Chamber 

m 


FIG.  166. — Longitudinal  Section  of  a  Puddling  Furnace. 


FIG.  167. — Sequence  of  the  Reactions  of  the  Puddling  Process. 

The  CO  coming  in  contact  with  the  air  burns  to  CC>2  with  its  characteristic 
blue  flame.  As  the  carbon  monoxide  increases  in  volume,  the  charge 
becomes  agitated  and  the  "  boil  "  is  in  progress.  The  charge  swells  and 
the  slag  pours  out  at  the  slit  beneath  the  working  door.  This  slag 
may  amount  to  12  to  25  per  cent  of  the  charge.  The  boil  continues  for 


320 


WROUGHT  IRON  AND  STEEL 


twenty  to  twenty-five  minutes  and  during  the  time  the  puddler  stirs  the 
charge  with  his  long-handled  rabble.  Toward  the  end  of  the  boil  the  metal 
begins  to  "  come  to  nature,"  and  pasty  masses  to  form  in  the  bath,  and  to 
show  above  the  slag.  Those  masses  are  stirred  and  gathered  by  the 
rabble  until  all  the  metal  becomes  pasty,  when  the  "  balling  "  period  begins. 
The  metal  at  this  time  is  gathered  into  three  or  four  portions,  each  of  which 
are  rolled  up  into  a  ball  made  up  of  many  particles  partly  welded  together. 
The  balls  are  rolled  up  near  the  bridge  out  of  the  flame  and  in  the  hottest 
place,  until  the  puddler  is  ready  to  draw  them.  They  are  removed  one  by 
one  for  squeezing,  after  which  the  hearth  is  repaired  for  the  next  charge. 

Squeezing  and  Working. — The  balls,  weighing  125  to  180  Ib.  each,  are 
withdrawn,  dripping  with  slag,  and  are  carried  to  the  jaws  of  a  squeezer 
by  which  most  of  the  slag  is  squeezed  out  and  then  made  smaller.  The 
squeezed  balls  are  sent  to  the  rolls  to  be  rolled  into  bars  called  "  Muck 
bar."  These  are  cut  into  lengths  and  wired  into  bundles,  half  the  bars 
piled  crossways  of  the  others.  These  bundles  are  reheated  in  a  reheating 
furnace  to  welding  heat  and  rolled  again  into  bars.  The  rerolled  material 
is  known  as  "  merchant  bar,"  and  the  effect  of  the  second  rolling  is  to 
eject  more  slag  and  to  form  a  cross-fiber  structure  as  the  result  of  the  cross 
piling. 

When  rolled  into  strips  this  is  called  "skelp."  Skelp  bent  into  shape  of 
tubes  and  butt-welded  or  lap-welded  makes  iron  pipe.  Steel  billets  are 
similarly  rolled,  forming  "  steel  skelp  "  and  made  into  steel  pipe. 

Following  is  the  composition  of  various  puddled  products : 


C. 

Si. 

s.     . 

p. 

Mn. 

Muck  bar  

0.10 

0.108 

0.052 

0.193 

Not  over  0  10 

Skelp  
Puddled  bar  
Wrought  iron  

0.05 
0.30 
0.10 

0.040 
0.120 

0.050 
0.134 

0.10 
0.139 
0.12 

Not  over  0  .  10 

Usually  there  is  more  than  1.0  per  cent  of  slag  in  wrought  iron  and  less 
than  0.2  per  cent  in  steel.  Ordinary  wrought  iron  is  practically  free  from 
manganese,  while  open-hearth  steel  will  contain  0.5  per  cent  or  more, 
hence  the  greater  liability  to  rusting  of  steel.  Most  of  the  wrought 
iron  made  in  the  United  States  goes  at  once  into  commerce;  a  little  is 
consumed  for  "  crucible  steel  "  (tool  steel). 


STEEL-MAKING 


In  the  year  1918,  of  the  39,000,000  tons  made  in  the  United  States, 
47  per  cent  was  for  basic  steel  and  33  per  cent  was  Bessemerized.  This  is 
due  largely  to  the  fact  that  a  pure  pig  iron  (Bessemer  pig)  is  needed  for 


BESSEMER  PROCESS  OF  STEEL  MAKING 


321 


this  process,  while  the  basic  open-hearth  can  handle  the  pig  from  impure 
ores  high  in  phosphorus,  which  are  more  plentiful.  Also  basic  open- 
hearth  steel  has  become  of  as  good  quality  as  Bessemer  steel.  The  duplex 
process,  later  described,  combines  the  speedy  steel-making  by  the  con- 
verter with  the  efficiency  of  the  open  hearth  in  purifying. 

STEEL-MAKING  BY  THE  ACID  BESSEMER  PROCESS 

The  pig  iron  used  in  the  Bessemer  process  preferably  contains  1  per 
cent  silicon  and  0.5  per  cent  manganese,  but  to  make  a  salable  steel,  the 


FIG.  168. — 400-ton  Hot-metal  Mixer. 

phosphorus  should  be  below  0.10  per  cent  and  the  sulphur  below  0.08  per 
cent,  since  neither  element  is  removed  in  the  converter.  If  the  silicon  is 
above  1  per  cent  the  large  quantity  of  slag  produced  carries  away  iron.  If 
far  below  1  per  cent,  the  charge  does  not  blow  hot.  When  manganese 
is  high  (1.5  per  cent),  it  makes  the  charge  sloppy,  the  slag  then  being  highly 
fluid  and  easily  ejected  during  the  blow. 

The  converters  in  a  large  plant  are  supplied  from  several  blast-furnaces, 
and  to  insure  a  good  average  pig  metal,  it  is  customary  to  collect  the  prod- 
uct of  the  several  furnaces  in  a  single  tilting  reverberatory  furnace,  or  hot 


322  WROUGHT  IRON  AND  STEEL 

metal  mixer  capable  of  holding  300  to  1300  tons  of  pig  metal.  From  the 
mixer  it  is  drawn  to  the  converters  as  needed,  and  a  regular  supply  is  thus 
assured. 

The  Hot-metal  Mixer. — Fig.  168  is  a  section  of  a  400-ton  mixer. 
Like  the  tilting  open-hearth,  this  is  carried  on  rollers  so  that  its  contents 
can  be  poured  into  a  casting  ladle  to  go  to  the  converters. 

It  is  driven  by  two  75  H.P.  electric  motors  which  act  in  series  during 
pouring,  but  in  parallel  during  the  return  of  the  mixer  to  its  normal  posi- 
tion. It  is  lined  with  13J  in.  of  firebrick  face,  in  order  to  cut  down 
radiation,  with  9  in.  of  magnesite  firebrick. 

The  Converter. — The  conversion  is  done  in  an  upright  converter,  lined 
with  silicious  material  held  together  with  fireclay.  Fig.  104  represents 
views  of  a  converter.  It  is  9  ft.  diameter  by  15  ft.  6  in.  high  and  is 
capable  of  treating  a  charge  of  20  tons  of  pig-metal.  It  is  swung  on 
trunnions,  through  one  of  which  the  compressed  air  needed  in  operation 
enters  to  the  tuyeres  at  the  bottom.  The  slag  made  in  a  converter  is  high 
in  silica,  and  has  but  little  effect  on  the  lining,  so  that  this  lasts  several 
months.  The  mouths  of  the  tuyeres  at  the  bottom  come  in  contact  with 
the  iron  oxide  formed  during  the  blow,  and  hence  this  part  of  the  con- 
verter lasts  only  twenty  to  twenty-five  hours.  This  bottom  accordingly  is 
made  so  that  it  can  be  replaced  by  another,  causing  a  delay  of  twenty  min- 
utes in  changing. 

Converter  Lining. — For  the  acid  process  the  converter  is  lined  to  the 
thickness  of  26  to  30  in.  with  ganister,  that  is,  quartz  rock  mixed  with  some 
clay  to  bind  it.  In  a  20-ton  converter,  Fig.  169,  there  are  fourteen  tuyeres 
of  well-burned  clay,  6  in.  diameter  by  30  in.  long,  each  tuyere  having  eight 
T^-in.  holes  extending  from  top  to  bottom  for  the  blast-air.  The  tuyeres 
are  set  in  place  upon  the  bottom  and  the  ganister  is  rammed  around  them. 
This  work  is  done  at  the  bottom  house  (see  the  general  plan  of  duplex  and 
electric  furnace  building),  where  are  situated  the  bottom-ovens  and  the 
grinding  and  mixing  machines  for  preparing  the  ganister. 

The  bottoms  on  transfer  trucks  are  run  into  ovens  15  ft.  square, 
where  they  are  thoroughly  dried  out.  These -are  heated  by  coal,  using  a 
forced  draft.  A  bottom  may  last  for  30  to  35  heats,  or  only  for  a 
single  one.  The  side-lining  is  of  ganister,  especially  around  the.  nose. 
Bottoms  are  changed  commonly  by  unbolting  while  the  vessel  is  bottom 
side  up,  then  lifting  it  off  by  crane. 

Operation  of  the  Acid-lined  Converter. — The  hot  converter,  from  which 
the  metal  of  a  blow  has  just  been  poured,  is  placed  in  a  horizontal  position 
and  15  tons  pig  iron  is  poured  into  it  by  means  of  a  ladle  that  is  brought 
from  the  mixer.  When  the  converter  is  in  this  position  no  metal  can  flow 
into  the  tuyeres  and  obstruct  them.  After  the  metal  is  poured  in,  the  blast 
(or  "  wind  ")  is  applied  at  the  rate  of  25,000  cu.  ft.  per  minute,  the  con- 


BESSEMER  CONVERTER 


323 


L 

o 

1 

324 


WROUGHT  IRON  AND  STfiEL 


verier  being  at  this  time  turned  to  the  vertical  position.  The  blast  now 
blows  in  fine  streams  upward  through  18  in.  of  molten  metal.  Active  oxida- 
tion of  the  manganese  and  silicon 
results  and  in  about  four  minutes 
they  are  oxidized  by  the  oxygen  of 
the  air  and  have  become  slag.  The 
carbon  now  begins  to  oxidize  to  CO, 
and  this  also  streams  upward  through 
the  metal  and  issues  with  the  air  from 
the  mouth  of  the  converter  in  a 
body  of  flame.  After  another  six 
minutes  the  flame  shortens  or  drops, 
and  the  operator,  knowing  that  the 
carbon  has  been  eliminated,  turns  the 
converter  into  horizontal  position,  the 
wind  being  at  the  same  time  shut  off. 
In  anticipation  of  this,  a  weighed 
quantity  of  spiegel  iron  or  "  Spiegel  " 
has  been  tapped  from  the  Spiegel-cupola,  where  it  is  kept  melted,  into  a 


FIG.  170.— Mixing  Pan,  Philips  & 
McLaren,  Pittsburgh. 


1703 

FIG.  171. — Worm-geared  Bottom-tap  Ladle, 
Pittsburg  Elect.  Furnace  Corp. 


FIG.  172.— Ingot  Mould. 


ladle.     The  ladle  is  transferred  by  the  traveling-crane  and  poured  into  the 
converter.     So  great  has  been  the  heat  evolved  by  the  oxidation  of  the 


ACID  BESSEMER  CONVERTING 


325 


impurities  of  the  pig  during  the  ten  minutes  of  the  blow  that  the  tempera- 
ture is  higher  than  at  the  start,  and  we  have  a  white-hot  liquid  consisting 
of  comparatively  pure  metal.  Oxidation-products  remain  hi  the  bath, 
and  the  carbon  and  manganese  of  the  charge  tend  to  reduce  these,  the 
unused  carbon  being  in  sufficient  quantity  to  impart  the  desired  strength 
to  the  steel.  Silicon,  which  also  is  introduced,  tends  to  dispose  of  gas 
contained  in  the  metal.  After  the  speigel  or  "  recarburizer  "  has  been 
added  and  the  reactions  have  ended,  the  steel  is  poured  from  the  converter 
into  the  ladle,  as  shown  at  the  left  of  Fig.  174,  also  the  ladle  is  hi  position  to 
receive  the  steel.  This,  after  a  short  interval,  is  carried  to  a  position  over 
the  ingot  molds  into  which  the  steel  is  to  be  teemed  or  poured.  The 
teeming-ladle,  Fig.  171,  is  "  bottom-poured,"  that  is,  a  tap-hole  and  plug 
are  arranged  in  the  bottom,  so  that  when  the  ladle  is  brought  over  the 
ingot  mold  a  stream  of  metal  drops  straight  downward  into  it  until  it  is 
filled;  in  this  way  the  molds 
are  filled  successively  until 
the  ladle  has  been  emptied. 

The  stopper-rod,  actuated 
by  a  lever-arm  outside,  plugs 
the  hole  from  within.  The 
end  of  the  rod  is  covered  by  a 
fireclay  lining,  to  withstand 
the  attack  of  the  hot  metal. 
The  ladle  is  tipped  by  operat- 
ing a  hand  wheel  or  by  aux- 
iliary hoist,  to  pour  out  the 
slag  that  remains. 

Fig.  172  is  the  ingot- 
mold  having  lugs  near  the 
top  by  which  the  mold  is 
picked  up  by  crane  when 
ready  for  stripping,  leaving 
the  ingot  standing  on  the 
car.  The  handle  on  the  side 
is  where  the  crane  takes  hold 
when  the  mold  is  to  be  laid 
on  its  side  for  cleaning. 

The  metal  remains  until 
cool,  after  which  the  molds 


2.005 


1.00^ 


o    i 


4       56       78      9 
Minutes  of  Blowing 


10     11     12   13 


FIG.  173. — Acid  Bessemer  Blow  (American 
practice). 


are  stripped  or  lifted  off,  leaving  the  ingots  standing.  The  ingot  is  picked 
up  and  conveyed  to  a  reheating  furnace,  and  finally  sent  to  the  rolls  to 
be  formed  into  the  shapes  desired  for  market  use.  Fig.  173  illustrates 
graphically  by  curves  the  progress  of  the  reactions,  and  the  elimination 


326  WROUGHT  IRON  AND  STEEL 

of  impurities  during  the  blow.  From  it  we  see  the  rate  at  which  the 
easily  oxidized  manganese  and  silicon  are  burned  and  also  the  carbon, 
which  is  but  little  acted  upon  until  these  disappear,  but  which  after  they 
are  gone  oxidizes  rapidly.  The  pig  contains  at  the  beginning  3.5  per 
cent  C,  1.0  per  cent  Si,  and  0.5  per  cent  Mn,  all  being  removed.  The 
recarburizer  adds  to  it,  as  Fig.  173  indicates,  1  per  cent  Mn,  0.7  per 
cent  C,  and  0.15  per  cent  Si.  The  manganese  is  added  to  take  from  the 
metal  the  oxygen  absorbed  during  the  blow;  the  carbon  is  to  give  the  steel 
the  required  strength  and  hardness,  and  the  silicon  to  dispose  of  the  gas 
contained  in  the  bath. 

THE  BASIC  BESSEMER  PROCESS 

The  converter  has  a  basic  lining  of  dolomite  mixed  with  tar  stamped  into 
place.  The  process  is  used  where  it  is  desired  to  treat  high  phosphorus 
ores.  It  is  little  used  in  the  United  States. 

A  typical  pig,  such  as  is  used  in  this  practice,  would  contain  C,  4.25 
per  cent;  Si,  1.0  per  cent;  S,  0.05  per  cent;  P,  1.5  per  cent;  Mn,  0.2  per 
cent.  This  is  charged,  using  11  tons  of  the  pig  with  the  addition  of  2600 
to  2800  Ib.  of  burned  lime  to  ensure  a  basic  slag.  The  blown  metal  would 
still  retain  C,  0.03  per  cent;  P,  0.07  per  cent;  S,  0.05  per  cent;  while  the 
slag  would  contain  SiO,  14  per  cent;  CaO,  48  to  51  per  cent;  MgO,  2  to  4 
per  cent;  P2Os,  17  to  19  per  cent;  Mno  and  FeO,  14  to  16  per  cent.  The 
high  phosphorus  content  makes  it  a  good  fertilizer  and  it  is  so  used. 

STEEL-MAKING  IN  THE  OPEN-HEARTH  FURNACE 

This  reverberatory  furnace  (Fig.  175)  is  used  in  the  melting  down  of  the 
materials  used  in  the  manufacture  of  steel. 

The  Two  Processes. — There  are  two  processes  of  producing  steel  in  the 
open-hearth  furnace,  called  respectively,  the  acid  and  the  basic.  The 
only  difference  in  the  open-hearth  furnace  used  is  that  for  the  acid 
process  the  hearth  is  lined  with  a  sand;  in  the  basic  process  with  basic 
material  such  as  dolomite  or  magnesite. 

The  Acid  Process. — By  the  acid  process  the  carbon,  silicon,  and  man- 
ganese,  impurities  of  the  molten  charge,  are  removed,  but  no  phosphorus 
or  sulphur  is  eliminated.  Hence  the  acid  open-hearth  can  only  treat  pure 
ores  of  the  Bessemer  type  low  in  these  two  latter  metalloids. 

The  Basic  Process. — On  the  other  hand  the  basic  process  can  remove 
from  the  charge  when  melting  not  only  carbon,  silicon,  and  manganese,  but 
also  sulphur  and  phosphorus.  Thus  a  charge  of  non-Bessemer  material 
can  be  used,  since  by  the  basic  process  it  is  possible  to  eliminate  the  sul- 
phur and  phosphorus  below  the  permissible  limit  of  0.05  per  cent  sulphur 
and  0.095  per  cent  phosphorus  necessary  for  a  good  grade  of  steel. 


OPEN-HEARTH  FURNACE 


327 


THE  OPEN-HEARTH  REVERBERATORY  FURNACE 

Fig.  175  is  a  perspective  view  of  the  front  of  a  stationary  open-hearth 
furnace  with  its  three  charging  doors.  In  183  is  a  transverse  section  of 
one  of  the  tilting  kind,  to  be  later  described.  The  furnace  is  fired  by  pro- 
ducer gas,  see  the  Hughes  producer,  Fig.  14. 

During  the  past  fifteen  years  the  Bessemer  process  has  been  gradually 
giving  way  to  the  basic  open-hearth  process,  due  to  the  fact  that  low  phos- 
phorus ore  is  being  exhausted.  It  is  claimed  that  for  most  purposes  open- 


FIG.  174. — Converter  and  Mixer  Building. 

hearth  steel  is  better  than  Bessemer,  but  the  latter  gives  the  most  satis- 
factory product  for  tin  plates,  and  is  well  suited  to  the  manufacture  of 
rails.  An  important  advantage  in  the  basic  open-hearth  process  is  that  it 
can  be  used  for  making  steel  from  pig  iron  and  ore  high  in  phosphorus. 

Fig.  176  is  a  half  sectional  plan  and  elevations  of  the  furnace,  having 
water-cooled  devices  designated  for  the  better  preservation  of  the  parts 
exposed  to  high  heat  and  corrosive  action.  It  is  basic-lined  with  material 
specified  by  the  legend  annexed. 

The  furnace  hearth   H  is  rectangular  and  open  at  each  end    o  t 
admission  of  air  and  gas  at  the  ports  C  and  D  respectively.     The  roof  is  of 
silica  brick,  12  in.  thick.     The  whole  furnace  is  heavily  ironed.     The  entire 


328 


WROUGHT  IRON  AND  STEEL 


bottom  and  hearth  of  the  furnace  is  built  in  and  supported  by  a  pan  of 
heavy  plates,  riveted  together  and  supported  on  I-beams  resting  on  piers. 
At  the  skew-backs  are  water-cooled  plates  set  against  the  I-beam  buck- 
staves  to  receive  the  thrust  of  the  arch.  On  the  charging  side  are  shown  the 
five  charging  doors,  counterbalanced  and  lifting  vertically,  large  enough 


q 

s. 

o 


to  enter  the  charging  boat  using  a  mechanical  charging-machine,  Fig. 
181.  At  the  middle  of  the  back  side  is  the  tap-hole  where  slag  and  metal 
are  drawn  off. 

Underground,  and  at  one  side  at  each  end,  are  two  checker  or  regen- 
erator chambers,  one  for  the  preheating  or  regeneration  of  the  air,  the 
other  for  the  gas.  The  arrangement  of  these  is  well  shown  in  section 


OPEN-HEARTH  FURNACE 


329 


plan  and  elevation,  in  Fig.  178.  The  checkers  are  built  up  of  9-in.  bricks 
into  a  series  of  pre-heated  flues  through  which  the  air  and  gas  pass  to  their 
respective  horizontal  flues  and  vertical  uptakes,  the  ports  delivering  to  the 
hearth  of  the  furnace. 


The  general  arrangement  of  the  furnace  and  its  accessories  is  given  in 
the  sectional  plan,  Fig.  178,  particularly  the  passages  and  flues  that  lead 
eventually  to  the  stack.  As  indicated  by  the  figure,  gas  from  the  main 
gas  flue  and  air  through  an  open  valve  at  the  left  are  traveling  along  their 
respective  gas  and  air  checker  chambers  and  entering  the  furnace  by  the 


330 


WROUGHT  IRON  AND  STEEL 


FIG.  177. — Open-hearth  Furnace  (transverse  section). 


Main  Gas  Flue 


§H| 

Vr 
G 

^teJ 

H 

ckJr 
A 

•  ~  -i!      \f 
i/b    ' 

l\*"-'  J 

1  *•--  1 

°1  i 

=  |Ko 

i\3j 

Center  Line  of  Furnace 
'6"Lengt'h  of  B 


|s                          •    U  U 

% 

9 

. 

HL 

Casting  Pit? 

Casting  Pit 

i 

FIG.  178. — Sectional  Plan  of  Open-hearth  Furnace. 


REVERSING  VALVES,  OPEN-HEARTH  FURNACES 


331 


gas  and  air  intakes  and  ports  at  the  left.  They  are  leaving  the  hearth  by 
the  uptakes  on  the  right,  passing  through  both  of  the  checker  chambers  and 
thence  through  a  three-way  valve  at  the  point  marked  "  close,"  "  open  " 
to  the  stack. 

The  flow  of  gases  having  passed  in  this  direction  for  fifteen  minutes, 


A*  'B   *  j-|    C 

FIG.  179. — Reversing  Valves  (gas  to  left  end  of  furnace). 


FIG.  180. — Reversing  Valves  (gas  to  right  end  of  furnace). 

the  valves  are  reversed  and  the  gases  pass  in  the  opposite  direction  for  a 
like  time.  During  the  fifteen  minutes  the  flame  is  drawn  through  the 
checker  chambers  highly  heating  them.  On  reversal  of  the  gas  current 
the  producer  gas  and  the  air,  before  they  reach  the  hearth,  are  heated  by 


332 


WROUGHT  IRON  AND  STEEL 


the  checker-work  and  both,  thus  preheated,  burn  at  a  higher  temperature. 
By  successive  reversals  the  temperature  of  the  regenerators  is  raised  and 
with  it  the  intensity  of  the  flame  of  the  air-gas  mixture.  This  accretion 
of  heat  can  go  on  to  the  point  where  the  brick  lining  begins  to  soften  and 
to  a  temperature  that  readily  melts  the  furnace  charge. 

Figs.  179  and  180  show  a  sectional  elevation  of  the  three-way  reversing 
valve,  already  referred  to.  In  the  upper  view  the  producer  gas  is  seen 
passing  through  a  disk- valve  D,  and  by  the  port  A,  to  the  furnace,  while, 
at  the  same  time  the  escaping  gases  passing  through  the  checker  cham- 
bers at  the  right  and  the  port  C,  go  through  B  to  the  chimney  or  stack  of 
Fig.  178.  The  valves  for  the  admission  of  air  and  of  gas  are  of  the  disk  type. 

Tapping. — At  Fig.  178  are  to  be  seen  the  two  casting  pits  where  the  ladles 
are  set,  adjoining  the  tap-hole  and  tapping-spout  S.  When  the  heat  is 
ready  to  tap,  a  bar  is  driven  through  the  tap-hole  and  enlarged  by  reaching 
through  from  the  charging  side  by  a  long  bar  until  the  slag  and  metal  flow 
out  freely.  The  tap-hole  should  be  carefully  cleaned  out  after  each  heat, 
then  replugged  with  clay. 

Fuels. — The  fuels  employed  for  the  open  hearth  are  natural  gas,  pro- 
ducer gas,  and  oil,  and  of  these  natural  gas  is  the  best  where  it  can  be  had, 
as  indicated  by  the  following  table : 


Constituent. 

Natural  Gas, 
Per  Cent. 

Producer  Gas, 
Per  Cent. 

Carbon  dioxide       ... 

0  40 

5  7 

Carbon  monoxide 

1  70 

22  0 

Oxygen  

1  80 

0  4 

Ethylene 

1  40 

0  6 

Ethane 

12  95 

0  0 

Methane  
Hydrogen    .        .    . 

68.85 

2.6 
10  5 

Nitrogen 

12  90 

58  2 

The  heating  power  of  natural  gas  is  550  CaL,  and  of  producer  gas  150 
Cal.  Oil  is  an  excellent  open-hearth  fuel.  It  can  be  vaporized  by  steam  or 
air  jet,  and  needs  no  preheating.  The  flame  of  it  is,  however,  sharp  and  is 
liable  to  cut  out  the  roof  and  to  over-oxidize  the  metal  of  the  molten  bath. 

The  Tilting  Open-hearth  Furnace. — This  differs  from  the  stationary 
type  chiefly  in  that  the  entire  furnace  body  may  be  tilted  or  rotated 
through  a  considerable  arc,  thus  pouring  slag  or  metal  at  any  stage  of  the 
process,  frequently  a  great  advantage. 

A  cross-section  of  such  a  furnace  is  shown  in  the  sectional  elevation  of 
an  open-hearth  building,  Fig.  183.  This  shows  the  furnace,  in  melting 
position.  This  type  of  furnace  does  away  with  tap-hole  troubles,  the 


ACID  OPEN-HEARTH  PROCESS 


333 


tap-hole  being  above  the  slag-line  in  melting  position;  also  the  furnace  can 
be  readily  emptied  between  heats  and  drained  easily  to  make  repairs. 

Mechanical  Charging. — At  181  is  a  perspective  view  of  a  charging 
machine.  In  this  view  is  shown  one  of  a  line  of  trucks  carrying  the 
charging  boxes.  These  are  picked  up,  one  at  a  time,  by  means  of  a  charg- 
ing ram,  thrust  through  the  charge-door,  inverted  to  discharge  their  con- 
tents of  steel  scrap,  then  withdrawn  and  set  once  more  on  the  truck.  In 
this  way  the  contents  of  box  after  box  is  put  into  the  furnace. 


FIG.  181. — Charging  Machine. 
THE  ACID  OPEN-HEARTH  PROCESS 

Object  to  be  Attained. — The  process  aims  to  reduce  within  defined 
limits  the  carbon,  silicon,  and  manganese  present  in  the  charge  of  scrap 
and  iron,  but  leaves  unchanged  whatever  sulphur  or  phosphorus  there  is. 

The  Charge. — This,  of  say  100,000  Ib.  weight,  is  made  up  commonly  of 
scrap  steel  and  pig  iron,  the  usual  average  being  50  to  75  per  cent  scrap, 
the  remainder  pig,  the  proportions  depending  on  supply  and  cost,  and 
being  such  as  to  produce  6  to  10  per  cent  of  slag  preferably  of  the  compo- 
sition 50  per  cent  silica  and  45  per  cent  of  FeO  and  MnO  together. 

The  following  are  representative  percentage  analyses: 


Elements. 

Pig-iron. 

Structural 
Steel  Scrap. 

Rail 
Steel  Scrap. 

Carbon 

3  00  to4  00 

0  20 

0  45 

Silicon  

1.00  to  2  00 

0  10 

0  15 

Manganese  
Phosphorus 

Under    1.00 
0  10 

0.50 
0  04 

0.90 
0  10 

Sulphur  

0.05 

0.04 

0  75 

334 


WROUGHT  IRON  AND  STEEL 


Computing  the  silicon  to  silica  and  the  manganese  to  MnO,  the  latter 
in  the  charge  should  be  less  than  half  the  silica.  Silica  comes  also  from  the 
sand  attached  to  ordinary  pig  iron  and  is  yielded  by  corrosion  of  the  sili- 
cious  bottom. 

Melting  Down. — By  the  time  the  charge  is  melted  down  both  man- 
ganese and  silicon  have  been  oxidized,  and  the  resultant  silica  has  united 
itself  to  the  MnO  and  the  little  iron  oxide  in  the  stock,  since  the  melting 
has  been  effected  with  a  natural  or  even  a  reducing  flame.  In  the  next 
stage,  in  a  hot  furnace  with  an  oxidizing  flame,  iron  is  oxidized  to  enter  the 
slag,  and  carbon  (till  then  but  little  affected)  is  burned  off.  To  aid  this 
operation  1000  to  2000  Ib.  of  iron  ore  is  added  in  calculated  proportions  and 
reacts  with  the  carbon  of  the  charge  thus : 


(13) 


Fe2O3+C  =  FeO+CO. 


The  iron  oxide  enters  the  slag,  the  CO  is  burned  to  CO2.  The  ore  is  added 
gradually  according  to  the  judgment  of  the  melter,  faster  in  a  hot  furnace 
and  according  to  the  character  of  the  slag. 

The  following  instructive  table  shows  the  composition  of  the  charge 
both,  before  and  after  melting  and  of  the  resultant  slag : 


Elements. 

GROUP  I. 
19  HEATS,  SOFT-COAL  GAS. 

GROUP  II. 
6  HEATS,  OIL  GAS. 

After  Melting, 
Per  Cent. 

End  of 
Operation, 
Per  Cent. 

After  Melting, 
Per  Cent. 

End  of 
Operation, 
Per  Cent. 

In 

Metal 

In 

Slag 

(Si.. 

0.02 

0.09 
0.54 
50.24 
21.67 
23.91 
45.58 

0.02 

0.04 
0.13 
49.40 
16.50 
29.79 
46.29 

0.5 
0.6 
0.64 
49.46 
13.16 
33.27 
46.43 

0.01 
0.02 
0.12 
49.36 
11.30 
34.11 
45.41 



Mn 

lc  

[SiO2 

MnO  

FeO  
MnO-FeO 

The  acid  open-hearth  process,  due  to  its  limitations,  is  decreasingly  in  use. 

THE  BASIC  OPEN-HEARTH  PROCESS 

To  make  steel  by  this  process,  lime  is  added  to  the  charge  to  produce  a 
basic  slag,  and  the  hearth  is  lined  with  basic  material  to  withstand  this 
basic  slag.  Iron  and  scrap  steel  that  contain  phosphorus  are  used.  There 
are  in  the  United  States  vast  bodies  of  non-Bessemer  ores  yielding  a  pig- 
iron  too  high  in  phosphorus  for  the  acid  open-hearth  process,  and  too  low 
for  the  basic  Bessemer  converter,  but  which  the  basic  open-hearth  can 
remove  without  difficulty  for  the  production  of  a  suitable  steel. 

The  method  employed  for  the  removal  of  carbon,  silicon,  and  manganese 


REACTIONS  BASIC  OPEN-HEARTH  PROCESS 


335 


are  the  same  as  in  the  acid  open-hearth  process,  except  that  in  basic  prac- 
tice there  is  an  addition  of  lime  for  the  formation  of  a  distinctly  basic  slag 
which  will  not  attack  the  basic-lined  bottom. 

Under  the  oxidizing  action  of  the  flame,  and  by  the  addition  of  some  iron 
ore,  the  phosphorus  is  oxidized  to  phosphoric  acid,  and  the  sulphur  is 
removed  as  calcium  sulphide  and  manganese  takes  up  a  farther  amount  as 
manganese  sulphide. 

The  phosphorus,  carbon,  silicon  manganese,  and  sulphur  are  eliminated 
by  oxidizing  them,  the  oxygen  being  obtained  principally  from  the  iron 
ore.  The  reactions  which  take  place  are  as  follows : 


(14) 
(15) 
(16) 

(17) 
(18) 


2P+5Fe2O3  =  P2O5+ lOFeO, 

C+FeO  =  CO+Fe, 
Si+2FeO  =  SiO2+2Fe, 

Mn+Fe203  =  MnO+2FeO, 
S+2FeO  =  SO2+2Fe. 


Of  the  products  of  these  reactions  the  CO  and  S02  are  volatile,  and  escape 
as  fast  as  formed,  the  speedy  escaping  of  the  CO  causing  the  boiling  of  the 
bath.  The  phosphoric  acid  (P2O5),  silica  (SiO2),  and  manganese  oxide 
(MnO)  separate  from  the  molten  iron  and  unite  with  any  bases  present  to 
form  the  slag,  a  phosphate  and  silicate  of  iron,  manganese,  lime,  magnesia 
and  alumina,  the  slag  floating  upon  the  surface  of  the  bath  to  be  removed 
by  pouring. 

CALCULATION  OF  CHARGE 
BASIC  OPEN-HEARTH  CHARGE-SHEET 


Weight 
Pounds. 

SiOi+PzOs. 

FeO  +MnO. 

CaO+MgO. 

Per  Cent. 

Pounds. 

Per  Cent. 

Pounds. 

Per  Cent. 

Pounds. 

Pig  iron  

50,000 
60,000 
1,500 
8,000 

4.0 

0.2 
5.0 
1.0 

2,000 
120 

100 
80 

1.0 
1.2 

77.0 
0.6 

500 
720 
1155 

48 

54.7 

4376 

4376 

Steel  scrap  

Iron  ore 

Limestone  

2300 

2423 

Slag 

SiO2+P2O5  =  23.0  per  cent 
FeO  +  MnO=23.0or  1  SiO2  =  l  FeO+MnO. 
CaO+MgO  =  46.0  or  1  SiO2  =  2  CaO+MgO 
Other  elements  8.0 


100.0 


336 


WROUGHT  IRON  AND  STEEL 


The  Charge. — Above  is  given  a  typical  charge  to  produce  a  slag 
of  the  composition  given.  The  proportions  of  pig  and  scrap  depend 
on  the  cost,  abundance,  and  relative  analyses  of  the  pig  and  scrap.  Both 
these  should  be  as  low  as  possible  in  sulphur  contents,  so  that  in  the  pigs, 
for  instance,  this  should  be  under  0.05  per  cent. 

Method  of  Charging. — The  common  method  is  to  charge  practically  all 
of  the  limestone,  then  the  pig  iron,  and  lastly  the  steel  scrap.  A  portion 
of  iron  ore  is  also  usually  charged  with  the  limestone  and  the  scrap,  the 
heat  then  has  the  benefit  of  the  oxidizing  action  all  the  time  that  the  metal 
is  in  the  furnace. 

Calculation  of  the  Charge. — Referring  to  this  charge  we  give  the  follow- 
ing analysis  of  its  constituents : 


Elements. 

Pig  Iron. 

Steel  Scrap. 

Iron  Ore. 

Limestone. 

Si(SiO2)  
P(P2OO                .    .  . 

0.75  (1.6) 
1.04  (2.4) 

Trace 
0.10  (0.2) 

(5.0) 
0.04  (0.08) 

(i.o) 

(0.01) 

Mn(MnO)  
Fe(FeO) 

0.75  (1.0) 

0.50(1.2) 

60  0  (77.0) 

0.6 

CaO              

0.2 

53.6 

MgO 

0.1 

1.1 

c        

4.00     . 

0.12 

s               

0.05 

0.06 

Trace 

Trace 

The  silicon  is  oxidized  to  SiCb,  the  phosphorus  to  P20s,  the  manganese 
to  MnO  by  the  air  and  the  iron  ore.  The  percentage  of  each  element  is 
multiplied  by  its  factor  to  express  its  amount  when  oxidized,  viz.,  SiO  =  2.1 ; 
Si,  P2O5  =  2.3  P;  MnO  =  1.3  Mn.  The  computations  are  quite  simple. 
It  will'  be  noted,  for  the  slag  specified  on  the  charge-sheet,  the  combined 
FeO-j-MnO  should  be  equal  to  the  combined  silica  and  phosphoric  acid, 
the  alkaline  bases  (CaO + MgO)  twice  that  quantity.  As  figured,  the  lime- 
stone might  be  increased  slightly. 

A  typical  open-hearth  charge  for  a  50-ton  furnace  is  as  follows:  Molten 
pig  iron  from  the  mixer,  50,000  lb.;  steel  scrap,  60,000  lb.;  limestone, 
8000  lb.  After  melting,  the  additions  in  the  furnace  would  be:  Iron  ore 
fed  in,  1500  lb.;  feldspar,  250  lb.  (to  promote  fluidity).  The  additions  in 
the  ladle  are  coke,  280  lb. ;  ferro-manganese,  500  lb. ;  aluminum,  1  lb.  The 
ordinary  method  is  to  charge  all  the  limestone,  then  the  molten  pig  and 
lastly  the  scrap. 

Operation. — As  shown  in  Fig.  182,  it  takes  four  hours  to  melt  a  charge, 
and  six  additional  hours  to  complete  the  manipulation,  so  that  in  ten  hours 
the  charge  is  ready  to  draw.  During  the  three-  to  four-hour  melting  period, 
the  carbon,  manganese,  and  silicon  we  can  see  are  reduced.  The  reactions 
are  controlled  by  the  melter,  who  sees  that  the  carbon  is  eliminated  last, 


REACTIONS,  BASIC  OPEN-HEARTH  PROCESS 


337 


and  if  it  is  oxidizing  too  fast  he  must  "  pig  up  "  the  charge  by  the  addition 
of  pig  iron  to  increase  the  carbon.  On  the  other  hand,  if  phosphorus  is 
oxidizing  too  fast,  the  oxidation  of  the  carbon  can  be  hastened  by  "  oreing 
down  "  (adding  iron  ore)  to  produce  the  following  reaction: 


(19) 


Fe2O3+3C  =  2Fe+3CO. 


If  carbon  is  eliminated  too  soon,  much  iron  becomes  oxidized.  With 
the  oxidation  of  silicon  and  phosphorus  to  silica  and  phosphoric  acid,  these 
acids  form  with  lime  and  iron  oxide  a  basic  slag  containing  10  to  20  per 
cent  SiO2,  5  to  15  per  cent  P2O5,  45  to  55  per  cent  CaO,  and  10  to  25  per 
cent  Fe.  The  slag  does  not  attack  the  basic-lined  hearth,  and  retains  the 
phosphorus  and  the  sulphur,  but  the  CaO  must  be  as  high  as  possible  for 


2.5 
2.0 

1.5 

•i.u 
OJ 


ft 


i 


V 


02  4  6  8  10 

Hours 

FIG.  182. — Chemical  Changes  in  the  Basic  Open-hearth  Furnace. 

this,  and  yet  not  so  high  as  to  render  the  slag  infusible.  After  melting, 
active  oxidation  begins,  and  the  bath  boils  by  the  escape  of  gas.  Upon 
the  completion  of  the  operation  the  charge  is  ready  for  tapping  into  a  50-ton 
ladle,  the  metal  filling  the  ladle  and  the  light  slag  overflowing  and  being 
thus  removed.  If  the  slag  remained,  phosphorus  would  be  reduced  from 
it,  upon  addition  of  the  recarburizer,  and  would  again  enter  the  steel. 

Recarburization. — Alike  in  acid  and  basic  practice,  this  signifies  the 
addition  of  ferro-manganese,  containing  both  manganese  and  carbon, 
which  restores  to  the  melted  charge  just  enough  of  these  elements  to  give 
the  desired  qualities  to  the  steel.  The  amount  needed  to  give  0.50  per 
cent  manganese  in  a  heat  of  100,000  Ib.  may  be  thus  calculated  for  Group  I, 
where  there  is  0.46  per  cent  or  460  Ib.  in  the  steel.  If  a  ferro  of  80  per  cent 
manganese  is  added  in  the  ladle,  when  the  steel  is  tapped  out,  there  will  be 
a  loss  of  25  per  cent  so  that  770  Ib.  of  ferro-manganese  should  be  added. 
Where  more  silicon  is  desired  it  can  be  supplied  by  the  use  of  ferro  silicon. 


338 


WROUGHT  IRON  AND  STEEL 


PQ 


OPEN-HEARTH  PLANT  339 

Alloy  steels  take  their  appropriate  metal.  To  increase  the  carbon  as  a 
hardener  charcoal  or  coke  in  paper  bags  is  thrown  into  the  ladle  and  half 
of  this  is  lost.  Aluminum  in  small  quantities  is  also  added  to  quiet  the 
metal  and  make  sound  ingots.  It  takes  the  oxygen  from  dissolved  iron- 
oxide  and  itself  rises  to  unite  with  the  slag  as  A^Os. 

Charge  Composition. — In  present  practice  the  charge  for  a  basic  fur- 
nace consists  of  steel  scrap  (steel  trimmed  in  the  process  of  manufacture, 
old  steel  rails,  and  steel  collected  by  junk  dealers) ;  of  pig-iron  containing 
less  than  1  per  cent  Si,  more  than  1  per  cent  Mn,  and  up  to  2  per  cent  in  P ; 


FIG.  184. — Crane  and  Magnet. 

of  calcined  limestone  (quicklime)  8  to  30  per  cent  of  the  charge;   and  of 
iron  ore. 

THE    OPEN-HEARTH   BUILDING 

Fig.  183  is  a  sectional  elevation.  Beginning  at  the  left  is  the  under- 
ground hopper,  vertical  elevator  and  storage  bin  for  the  coal  for  the  pro- 
ducers that  make  the  gas  for  the  three  200-ton  tilting  open-hearth  fur- 
naces. By  a  goose-neck  pipe  the  gases  pass  to  the  underground  system 
of  flues  and  chambers  as  already  described. 

On  the  elevated  track,  just  within  the  main  building,  stands  the  loco- 
motive which  brings  in  the  65-ton  metal  ladles.  These  are  picked  off 
their  trucks  by  the  100-ton  crane  and  poured  into  the  tilting  200-ton  open- 
hearth  furnace  as  shown.  Just  beneath  the  ladle  on  the  elevated  working 
platform  is  a  track  on  which  are  brought  in  the  boxes  of  scrap  or  pig 
needed  for  the  charge.  The  open-hearth  metal  when  finished  is  poured  into 
a  110-ton  bottom  tap  ladle,  and  then  picked  up  by  the  heavy  175-ton 
traveling  crane  by  which  it  is  tapped  into  the  ingot-molds  near  the  right 
side  of  the  open-hearth  building.  One  notes  the  small  recarburizing 
ladle  by  which  the  ferro-silicon,  etc.  (first  melted)  and  other  additions  are 


340  WROUGHT  IRON  AND  STEEL 

made.  The  supernatant  slag,  as  it  accumulates,  is  poured  into  a  slag 
pot,  set  directly  beneath  the  spout. 

The  three  open-hearth  furnaces  each  of  which  has  a  hearth  area  of  900 
sq.  ft.  are  electrically  operated.  They  are  so  constructed  as  to  be  heated 
either  by  producer  gas  or  by  fuel  oil.  Fig.  181  shows  the  charging 
machine.  The  cars  carrying  the  boxes  of  cold  stock,  whether  scrap  iron, 
steel,  or  pig  iron  are  set  in  front  of  the  furnace  charging-door.  The 
charging  bar  of  the  machine  hooks  upon  a  box,  lifts  it  off  the  transfer  car 
and  carries  into  the  furnace.  The  box  makes  a  half-turn  which  dumps 
the  load,  and  the  empty  box  is  at  once  withdrawn.  In  this  figure  the 
fixed  open-hearth  is  shown  as  in  Fig.  176:  in  Fig.  183  is  a  50-ton  tilting 
furnace.  Beneath  the  charging  floor  will  be  seen  a  section  of  the  gas  cham- 
ber leading  to  the  stack. 

The  casting  ladles  receive  the  finished  charge  or  heat  which  is  tapped 
into  ingot  molds  standing  on  the  pit  floor.  The  pit  slag  from  the  furnaces 
is  handled  in  steam-dumping  ladles. 

The  Gas-producer  Building  and  Stock  Yard. — This  building  (not 
shown)  is  parallel  to  the  open-hearth  building.  It  contains  nine  self- 
cleaning  Hughes  producers  (see  Fig.  14).  For  furnishing  basic  lining 
there  are  also  two  dolomite  kilns  and  a  crusher.  The  stock  yard,  where  is 
assembled  the  iron  and  steel  scrap  and  the  iron  ore  and  limestone,  is  located 
between  the  gas-producer  building  and  the  open-hearth  building.  In 
the  stock  yards  by  means  of  a  10-ton  magnet  crane,  Fig.  184,  the  scrap 
is  picked  up  and  loaded  into  the  charging  boxes.  The  boxes  are  then 
brought  into  the  open-hearth  building  and  placed  close  to  the  furnaces 
ready  to  be  loaded  into  any  one  of  them  by  means  of  the  charging  machine 
(see  Fig.  181). 

THE  DUPLEX  PROCESS  OF  STEEL  MAKING 

The  duplex  process,  now  largely  in  use  in  large  plants,  is  generally 
understood  to  mean  the  making  of  steel  from  non-Bessemer  pig  iron  by  a 
combination  of  the  acid-Bessemer  and  the  basic  open-hearth  process. 
The  acid-lined  converter  oxidizes  the  silicon,  the  manganese  and  a  certain 
portion  of  the  carbon  of  the  pig,  the  amount  of  the  carbon  depending^upon 
the  practice.  The  blown  metal  is  then  transferred  by  ladle  to  the  basic 
open-hearth  furnace  where  the  phosphorus  and  the  remainder  of  the  car- 
bon are  removed.  The  oxidation  or  burning  off  of  the  silicon,  manganese 
and  carbon  proceeds  rapidly  in  the  converter,  while  in  the  open-hearth 
the  phosphorus  as  it  oxidizes  enters  the  basic  slag  which  does  not  attack  a 
basic  lining.  The  duplex  process  shortens  the  open-hearth  purification 
by  more  than  five-sixths  of  the  usual  period,  giving  a  steel  of  the  same 
quality  as  the  straight  open-hearth  process. 


OPEN-HEARTH  AND  ELECTRIC  FURNACE  PLANT  341 

In  practice  the  pig-iron  is  poured  into  the  converter,  the  blast  turned 
on  and  the  heat  blown  until  in  the  judgment  of  the  blower  the  metal  is  of 
the  desired  metalloid  content.  In  the  case  that  high-phosphorus  iron  is 
used  the  blow  is  stopped  when  the  metal  still  retains  about  1.00  per  cent 
of  carbon ;  while,  when  treating  a  low  phosphorus  pig,  the  metal  is  nearly 
decarbonized.  The  blown-metal  together  with  2  to  3  per  cent  of  lime  to 
give  a  basic  slag  is  now  charged  to  the  open-hearth  furnace.  If  the  metal 
has  been  decarbonized  10  per  cent  molten  pig  iron  is  added  either  in  the 
transfer  ladle  or  in  the  open-hearth  furnace.  In  the  furnace  a  reaction 
takes  place;  the  phosphorus  oxidizes  and  enters  the  slag  as  phosphate  of 
lime  while  the  carbon  is  removed  as  carbon  dioxide.  When  the  phos- 
phorus is  within  the  specified  limits,  as  determined  by  a  rapid  laboratory 
analysis,  the  heat  (the  charge)  is  tapped  into  a  ladle  where  proper  addi- 
tion for  the  required  manganese  and  carbon  content  are  made  in  the  steel 
ladle.  The  usual  way  is  to  have  three  20-ton  converters  supplying  one 
sixty-ton  open-hearth  furnace  with  metal.  If  the  three  20-ton  converters 
are  blown  together  and  their  united  metal  assembled  in  one  transfer 
ladle  for  removal  to  the  open-hearth,  such  a  plant  can  keep  four  or  five 
open-hearth  furnaces  in  continuous  operation.  This  is  due  to  the  time 
needed  for  the  respective  purifications,  the  Bessemer  taking  from  fifteen 
to  twenty  minutes  while  the  open-hearth  will  take  from  ninety  to  110 
minutes. 

DUPLEX  AND  ELECTRIC-FURNACE  PLANT 

Fig.  185  gives  the  general  arrangement  of  the  converter  and  open- 
hearth  departments  of  a  large  plant  using  the  duplex  process  with  the 
addition  of  an  electric  furnace  building  and  a  forge-press  building.  It 
is  arranged  to  provide  Bessemer  metal  for  open-hearth  refining,  and 
open-hearth  metal  for  electric  refining;  also  directly  made  Bessemer 
ingots  and  open-hearth  ingots. 

The  location  of  the  Bessemer,  the  open-hearth  and  the*electric  furnaces 
is  shown,  also  the  bottom  house  where  the  converter  bottoms  are  made 
and  dried.  The  position  of  the  forge  press  building  is  also  indicated,  but 
the  latter  is  not  described. 

The  Converter  and  Mixer  Building,  Fig.  174,  shows  many  details 
of  operation  in  making  steel  from  Bessemer  pig  from  the  receipt  of  the 
molten  metal  from  the  blast-furnaces  to  the  production  of  the  ingot. 

The  molten  metal  is  tapped  from  the  blast-furnaces  into  a  "  65-ton 
ladle  "  carried  on  a  truck  at  the  ground  level  near  the  converter.  It  is 
lifted  from  the  truck  by  the  "  100-ton  crane/'  and  poured  into  the  "  1300- 
ton  mixer  "  where  a  large  body  of  molten  metal  is  accumulated  of  the 
average  grade  produced  by  the  blast-furnaces.  When  a  charge  is  needed 
for  any  converter,  the  mixer  is  tilted  and  a  part  of  its  contents  poured  into 


342 


WROUGHT  IRON  AND  STEEL 


ELECTRIC  STEEL  MAKING  343 

the  "  hot-metal  transfer  car  "  which  travels  on  an  elevated  platform  at 
the  height  needed  to  enable  it  to  pour  into  any  converter  when  this  is 
turned  down  into  receiving  position.  To  do  this,  the  ladle  being  in  posi- 
tion, a  hook  beside  it  tips  the  ladle.  The  hook  is  attached  to  a  steel  rope 
traveling  over  pulleys  and  actuated  at  the  converter,  so  that  as  the  con- 
verter turns  down  the  ladle  begins  to  pour.  As  the  converter  is  turned 
back  blowing  begins.  Additions  of  scrap  or  pig  metal  are  made  to  the 
mixer  or  to  the  converter  during  blowing  from  a  concrete  scrapping  plat- 
form above  them.  The  scrap  is  brought  in  boxes  carried  on  trucks 
and  these  boxes  are  hoisted  from  the  trucks  and  delivered  to  the  platform 
to  be  added  to  the  charge  as  needed. 

The  blown  metal  is  poured  from  the  converters  into  a  bottom-tap 
ladle  which  is  then  brought  over  the  ingot  molds  standing  upon  trucks  seen 
close  to  the  side  of  the  building  at  the  left.  The  ladle  is  bottom-poured  or 
teemed  into  the  molds.  For  closer  observation  in  pouring  the  craneman's 
cab  is  carried  on  a  frame  secured  and  braced  to  the  traveling  crane. 
A  platform  at  the  level  of  the  top  of  the  molds  is  for  the  tapper  who  does 
the  pouring.  Just  outside  the  building  and  near  by  is  the  blowing  plat- 
form where  the  converter  man  stands  to  operate  the  converter  80  ft.  away. 
The  air  supply  comes  from  a  pressure  blower  in  an  adjacent  building. 
When  cool,  the  ingot  trucks  are  taken  to  the  stripper  yard,  when  the 
molds  are  stripped  or  lifted  off  from  the  ingots  and  these  are  sent  to  the 
rolling  mill. 

ELECTRIC  STEEL-MAKING 

The  manufacture  of  electric  steel  is  becoming  well  established  in  the 
United  States,  the  estimated  tonnage  for  the  year  1919  being  1,215,000 
short  tons.  This  increase  has  been  due : 

(1)  To  the  production  of  a  more  uniform  quality  of  steel. 

(2)  To  the  fact  that  electric  steel  can  be  poured  at  a  much  higher  tem- 
perature when  still  or  dead,  ensuring  the  production  of  thinner  castings. 

(3)  That  the  tensile  strength  and  other  physical  properties  show  that 
the  steel  is  stronger  and  tougher  than  other  steel. 

In  addition  one  has  to  remember  that  in  the  United  States  in  1919 
there  were  about  85  electric  furnaces  in  the  non-ferrous  metal  trades  and 
about  100  producing  ferro-alloys. 

The  Electric  Furnace. — Figs.  186  and  187  are  perspective  views  of 
an  electric  furnace  and  control  panel  and  of  the  transformer  and  sub- 
station equipment  respectively.  As  shown  in  the  sectional  elevation, 
188,  it  is  a  three-electrode  tilting  furnace,  resembling  in  its  action 
a  great  arc-light.  It  uses  an  alternating  current.  The  current  in  the  sub- 
station is  transformed  from  the  supply  line  to  a  low-voltage,  high-amper- 
age current,  thus  giving  a  heavy  current  for  the  melting.  In  the  figure  a 


344 


WROUGHT  IRON  AND  STEEL 


clean  pit  is  shown  beneath  the  furnace  while,  in  Fig.  189,  the  furnace  is 
set  so  as  readily  to  pour  to  ladles  set  on  the  floor.  The  furnace  shell  is  a 
steel  pan  having  a  brick  and  ganister  lining  and  with  a  firebrick  roof.  It 


FIG.   186. — Electric-furnace  and 
Control-panel. 


FIG.  187. — Transformer  and  Sub- 
station Equipment. 


is  charged  with  steel  scrap  through  a  door  at  the  right,  and  the  electrodes 
are  lowered  upon  the  charge  heating  and  melting  it.  When  ready,  the  slag 
is  poured  off  through  a  spout  at  this  side,  while  the  metal  is  received  into  a 
ladle  at  the  left,  as  shown  in  Fig.  189. 


Column  Side 


FIG.  188. — Section  of  Electric-furnace  Showing  Lining  and  Bottom  Neutral 

Connection. 

The  Electric-furnace  Building. — Centrally  in  the  sectional  elevation  of 
Fig.  189  is  placed  the  three-pole  electric  furnace,  having  at  the  ground 
level  at  the  right  a  50-ton  steel  transfer  ladle  and  on  the  left  a  slag  pot  or 
car  for  removal  of  the  slag  from  the  furnace.  The  whole  building  is 


ELECTRIC-FURNACE  PLANT 


345 


commanded  by  a  60-ton  traveling  crane  for  charging,  and  removal  of  the 
finished  metal  in  a  30-ton  ladle  to  the  ingot  molds.  At  the  left  side  is  a 
5-ton  wall  crane  for  stripping.  The  mechanism  beneath  the  crane  is  for 
electrically  tipping  it  through  medium  of  a  sector  and  spur  gearing.  The 
furnace  has  three  carbon  poles,  as  shown  in  Fig.  188. 


FIG.  189. — Electric-furnace  Building. 

VARIETIES  OF  STEEL 

Basic  Open-hearth  Steel.— This  is  a  quite  pure  steel  containing  less 
than  0.10  per  cent  impurities. 

High-Grade  Steel. — A  steel  made  in  the  electric  furnace. 
Steel  rails  are  designated  as  follows: 


(Chemical  Specification,  Asso.  of  Amer.  Steel  Manufacturers) 


C. 

Si 

p. 

Mn. 

Bessemer 

0  35  to  0.55 

Not  over  0.20 

Not  over  0.10 

0  70  to  1  14 

Open  hearth  

0.46  to  0.75 

Not  over  0.20 

Not  over  0  .  04 

0.60  to  0  90 

346 


WROUGHT  IRON  AND  STEEL 


STEEL-CASTINGS  HAVE  THE  FOLLOWING  PROPERTIES 

(Extract  from  Specifications,  Amer.  Society  for  Testing  Materials) 


MINIMUM  PHYSICAL  REQUIREMENTS. 

MAXIMUM. 

Tensile 
Strength 
Lb.  per 
Sq.  In. 

Yield 
Point 
Lb.  per 
Sq.  In. 

Per 
Cent. 
Elong. 
in  2  In. 

Per 
Cent. 
Red. 
in  Area. 

C. 

s. 

p. 

Ordinary  castings  
Tested  castings,  hard  
Tested  castings,  medium.  .  . 
Tested  castings,  soft  

85,000 
70,000 
60,000 

N< 
38,250 
31,500 
27,000 

3ne  requir 
15 
18 
22 

ed 
20 
25 
30 

0.40 

0.05 
0.05 
0.05 

0.08 
0.05 
0.05 
0.05 

STRUCTURAL  STEELS  FOR  BUILDINGS  ARE  THUS  DESIGNATED 


Structural  Steel. 

Rivet  Steel  O.  H. 

Phosphorus,  maximum,  Bessemer 

0  10  per  cent 

Phosphorus,  maximum,  open  hearth  

0  06  per  cent 

0  06  per  cent 

Ultimate  tensile  strength,  pounds  per  square  inch  .... 
Yield  point 

55,000-65,000 
\  Ult  tens  str 

48,000-58,000 
i  Ult  tens  str 

Character  of  fracture 

Silky 

Silky 

Cold  bend  without  fracture  

180°  to  diam. 
of  1  thickness 

180°  flat 

TOOL  STEEL  IS  OF  THE  FOLLOWING  COMPOSITION 


Tung- 
sten. 

Chro- 
mium. 

Car. 

Sul. 

Phos. 

Sil. 

Vana- 
dium. 

Carbon  steel 

1  10 

0  03 

0  015 

0  20 

High-speed  steel  

18.00 

"  3.50 

0.55 

0  .  012  or  less 

Trace 

Trace 

1.00 

Alloy  steel  may  be  defined  as  ordinary  properly  melted  carbon  steel  to 
which  have  been  added  ferro  compounds  of  certain  rare  metals  in  sufficient 
though  small  amount  to  materially  modify  the  qualities  of  the  original 
carbon  steel  as  shown  in  the  following  table. 

The  alloy  steels  are  divided  into  two  groups,  those  with  one  metal  alloyed 
as  in  (3)  and  quaternary  steels  with  two  metals  alloyed  as  in  (7) .  They 
have  high  elastic  limit,  great  strength  and  toughness.  The  first  two  qual- 
ities are  enormously  increased  by  heat  treatment  (quenching  and  temper- 
ing) and  the  steel  still  retains  great  toughness.  The  most  important  of  the 
structural  alloy  steels  are  those  of  nickel,  chromium,  and  vanadium,  which 
by  heat-treatment  can  be  given  a  tremendous  range  of  strength,  varying 
from  100,000  to  250,000  Ib.  per  square  inch. 


STEEL  AND  IRON  PRICES 


347 


COMPOSITION  OF  ALLOY  STEELS 


Carbon 
Per  Cent. 

Manga- 
nese 
Per  Cent. 

Nickel 
Per  Cent. 

Chromium 
Per  Cent. 

Vanadium 
Per  Cent. 

Elastic 
Limit 
Lbs.  per 
Sq.  In. 

Tensile 
Strength 
Lbs.  per 
Sq.  In. 

Elonga- 
tion in  2 
In.  Per 
Cent. 

Reduc- 
tion of 
Area 
Per 
Cent. 

(1) 

27 

55 

49000 

80000 

30 

65 

(2) 

(3) 

.27 

45 

.47 
50 

.26 

66,000 
65,000 

98,000 
96,000 

25 
22 

52 

52 

(4) 

43 

60 

32 

96,000 

122,000 

21 

52 

(5) 

.30 

.60 

3.40 

75,000 

105,000 

25 

67 

(6) 

(7) 

.33 
30 

.63 
49 

3.60 

3  60 

1   70 

.25 

118,000 
119,000 

142,000 
149,500 

17 
21 

57 
60 

(8) 

.25 

.50 

2.00 

1.00 

102,000 

124,000 

25 

70 

(9) 

.38 

.30 

2.08 

1.16 

120,000 

134,000 

20 

57 

(10) 

.42 

.22 

2.14 

1  27 

.26 

145,000 

161,500 

16 

53 

(11) 

.36 

.50 

1.30 

.75 

.16 

140,000 

157,500 

17 

54 

(12) 

.30 

.50 

.80 



90,000 

105,000 

20 

50 

(13) 
(13) 

.23 
35 

.58 
64 

.82 
1  03 

.17 

22 

106,000 
132,500 

124,000 
149,500 

21 
16 

66 
54 

(15) 

.50 

.92 

1.02 

.20 

170,000 

186,000 

15 

45 

Lathe  tools  made  from  high-speed  steel  can  be  run  at  a  speed  of  30  ft. 
per  minute,  indeed  so  that  the  cutting  edge  shows  a  just  visible  red,  a 
speed  four  times  as  great  as  that  which  ordinary  tool  steel  will  stand. 


IRON  ORE  AND  PIG-IRON  PRICES 

Pig-iron,  Pittsburg  Market  in  1920  (Quotations  for  Carload  Lots).— 
Standard  Bessemer,  $29.35;  malleable  Bessemer,  $28.65;  basic,  $27.15; 
No.  2  foundry,  $28.15;  gray  forge,  $27.15.  Standard  Bessemer  is  used  for 
making  steel  in  the  Bessemer  converter,  malleable  Bessemer  for  malleable 
iron  castings,  basic  for  steel  suited  to  the  basic  open-hearth  furnace,  foundry 
for  making  foundry  castings  suited  to  machining,  and  gray  forge  for 
wrought-iron.  In  the  Chicago  markets  both  Southern  and  Northern 
pig-iron  are  quoted.  The  first,  from  the  great  iron  center  at  Birmingham, 
Ala.,  though  cheap,  is  high  in  phosphorus.  The  Northern  iron  from  nearby 
points  is  made  from  Lake  Superior  ores.  Pig  iron,  cast  in  sand,  is  weighed 
to  2260  Ib.  for  a  long  ton,  the  20  Ib.  excess  being  allowance  for  the  sand  that 
sticks  to  the  pigs. 

Steel,  Pittsburg  Market. — Bessemer  and  open-hearth  billets  are  quoted 
at  $38.50.  These  are  ingots  4  in.  square  by  6  ft.  long  that  are  re-heated 
and  rolled  into  the  required  merchant-steel  bars. 

Iron  Ores  are  purchased  by  guarantee  on  the  part  of  the  shipper  that 
they  will  come  up  to  a  given  standard  that  generally  is  based  upon  the 
percentage-content  in  natural  condition,  thus  including  the  contained 


348  WROUGHT  IRON  AND  STEEL 

moisture.     For  Lake  Superior  ore  the  prices  for  1920  at  Lake  Erie  ports, 
per  long  ton  (2240  Ib.)  were: 

Old  range  Bessemer,  55  per  cent  iron  base $6 . 45 

Old  range  non-Bessemer,  51.5  per  cent  iron  base 5 . 70 

Mesabi  Bessemer,  55  per  cent  iron  base 6 . 20 

Mesabi  non-Bessemer,  51.5  per  cent  iron  base 5 . 55 

On  a  non-Bessemer  ore,  as  it  varies  from  this,  a  premium  of  11.95  cents 
is  paid  for  each  per  cent  iron  over  the  guarantee,  and  a  penalty  or  reduc- 
tion is  made  of  11.95  cents  down  to  50  per  cent  iron,  and  17.95  cents  down 
to  49  per  cent  iron,  and  a  double  penalty  down  to  48  per  cent  iron.  Below 
48  per  cent,  the  penalty  becomes  27  cents  per  unit.  On  Bessemer  ore, 
provision  is  made  for  a  premium  only  in  case  the  ore  exceeds  the  guaran- 
teed 55  per  cent  iron. 

The  Old  Range  ores  come  from  the  iron  ranges  on  the  south  side  of 
Lake  Superior,  and  command  a  higher  price  because  of  the  better  mechan- 
ical condition.  The  Mesabi  ores  are  soft,  friable,  and  carry  much  fine, 
which  makes  flue-dust.  When  smelting  such  ore,  in  ratio  of  85  per  cent 
soft  to  15  per  cent  hard  ore,  as  much  as  6  per  cent  flue-dust  or  dirt  is  made. 

Steel  Works,  Scrap,  1920. — Heavy  melting  steel  per  gross  ton  delivered 
at  works  $19.00  to  $21.00. 

Cost  of  Production  of  Pig-iron  in  the  electric  furnace  in  1915  was 
$26.21  per  long  ton  and  of  steel  from  steel  scrap  $29.90  per  long  ton  based 
on  a  cost  for  common  labor  of  $2.50  per  eight-hour  shift. 


PART  V 
COPPER 


CHAPTER  XXVII 
COPPER  ORES  AND  THEIR  TREATMENT 

CHARACTERISTICS  OF  COPPER  ORES 

We  are  to  think  of  copper  ores  as  mineral  aggregates,  carrying  frequently 
10  to  15  per  cent  copper  or  less,  with  associated  minerals  and  an  earthy 
gangue.  Treating  the  ore  is  a  problem  not  only  of  obtaining  the  copper, 
but  of  separating  and  eliminating  the  gangue  and  associated  minerals. 
Though  there  are  many  kinds  of  copper  ores,  those  of  commercial  impor- 
Jance  are  few  in  number.  We  may  divide  them  into  three  classes:  (1) 
the  sulphides;  (2)  the  oxides,  including  the  carbonates  and  silicates;  (3) 
ores  containing  native  copper. 

Sulphides. — Chalcopyrite,  CuFeS2,  when  pure  contains  34.5  per  cent 
copper.  This  is  by  far  the  most  widely  distributed  and  most  abundant 
of  the  ores  of  copper,  and  furnishes  the  world's  principal  supply  of  the 
metal.  It  is  frequently  accompanied  by  iron  pyrite,  and  has  silicious 
gangue  even  when  the  sulphide  is  massive.  In  consequence,  the  ore  often 
carries  no  more  than  3  to  4  per  cent  copper,  but  is  particularly  suited  to 
pyrite  smelting,  that  is  to  smelting  with  little  fuel.  Silver  and  gold  are 
found  in  the  ore  in  small  quantity.  The  deposits  at  Mt.  Lyell,  Tasmania, 
that  have  been  so  successfully  worked  by  pyritic  smelting,  are  chiefly  of 
massive  iron  pyrite,  containing  chalcopyrite,  and  carrying  4.5  to  5  per  cent 
copper  with  0.15  oz.  gold  and  3  oz.  silver  per  ton. 

Chalcocite  (copper  glance),  CU2S,  is  computed  to  contain  79.7  per  cent 
copper,  but  it  is  seldom  pure  even  in  the  crystalline  form,  the  copper 
having  been  replaced  by  iron  and  other  metals.  The  impure  mineral 
shows  the  characteristics  of  the  pure  mineral  when  carrying  as  little  as 
55  per  cent  copper.  The  pure  crystals  resemble  the  artificial  product 
"  white  metal,"  a  high-grade  copper  matte  produced  in  the  furnace. 

Bornite  (peacock  ore),  CuaFeSs,  when  pure  contains  55.6  per  cent  cop- 
per. It  is  found  associated  with  chalcopyrite  and  chalcocite  in  propor- 
tions varying  from  42  to  70  per  cent  copper,  without  losing  the  character- 
istic varied  colors. 

Enargite  (3Cu2S-As2Ss),  48.3  per  cent  copper,  an  arsenide,  occurs  in 
Butte,  Mont.,  ores. 

Tetrahedrite      (gray     copper     fahlerz),  _  (Cu2S,FeS,ZnS,Ag2S,PbS), 

351 


352  COPPER  ORES  AND  THEIR  TREATMENT 


,  may  be  computed  as  containing  30.4  per  cent  copper,  but  it 
varies  greatly  in  the  copper  and  silver  content.  It  has  already  been 
mentioned  as  a  silver  ore.  Because  of  the  contained  arsenic  and  antimony 
it  is  unfavorable  as  a  copper  ore,  and  it  is  only  because  of  the  richness  in 
silver  that  it  is  treated. 

Oxides,  Carbonates,  and  Silicates.  —  These  ores  are  the  result  of  the 
decomposition  of  the  copper  sulphides,  by  air  and  water.  We  find  them 
in  the  upper  zones  of  mineral  deposits  accompanied  by  iron  oxide,  which 
also  is  the  result  of  the  decomposition  of  iron  sulphide.  As  we  sink  on 
the  vein  we  find  the  oxidized  ore  of  the  upper  levels  giving  place  in  depth 
to  the  unaltered  sulphides. 

Cuprite  (red  copper-oxide),  Cu20,  88.8  per  cent  copper,  is  a  product  of 
decomposition.  It  often  permeates  large  masses  of  iron  ore.  Large  lumps 
of  the  ore  are  sometimes  found,  the  center  of  which  contains  unaltered 
metal.  These  evidently  are  the  result  of  the  oxidation  of  a  mass  of  native 
copper. 

Melaconite  (black  oxide  of  copper),  CuO,  contains  when  pure  79.8 
per  cent  copper.  The  ore,  with  the  copper  in  part  replaced  by  oxides  of 
iron  and  manganese,  is  sometimes  found  in  masses  large  enough  to  pay  for 
extraction,  and  containing  20  to  50  per  cent  copper.  The  so-called  black 
oxide  of  the  Blue  Ridge  region,  on  the  border  of  Tennessee,  North  Carolina, 
and  Virginia,  seems  to  be  an  intimate  mixture  of  copper  glance,  black  cop- 
per oxide,  copper  carbonate,  and  native  copper  with  iron  oxide  and  sul- 
phide. The  ore  can  be  readily  roasted  in  lump  form. 

Malachite,  CuCC>3-Cu(OH)2,  57.3  per  cent  copper,  occurs  widely 
distributed,  ordinarily  in  non-paying  quantities  as  a  decomposition  prod- 
uct in  surface  deposits,  but  sometimes  sufficiently  rich  to  work.  It  is 
found  mixed  with  limestone,  dolomite,  oxides  of  iron,  manganese,  and 
silica.  It  is  difficult  to  judge  the  copper  content  of  the  ore  from  the 
appearance,  but  the  green  color  makes  its  presence  readily  recognizable. 

Azurite,  2CuCO3-Cu(OH)2,  is  computed  to  contain  55.2  per  cent  cop- 
per. The  ore  is  blue,  as  the  name  indicates,  and  the  appearance  is  striking. 
It  occurs  in  the  same  way  as  malachite,  and  often  is  associated  with  mala- 
chite, but  it  is  less  abundant,  and  often  is  only  a  coloring  on  other  oxides. 

Chrysocolla,  a  hydrated  silicate  of  copper,  containing  when  pure  40 
per  cent  copper,  is  a  decomposition  product  of  copper  sulphide,  and  is 
often  accompanied  by  malachite. 

Native  Copper.  —  Native  copper  is  found  extensively  in  the  copper  region 
of  Lake  Superior.  Elsewhere  it  occurs  sparingly  and  is  not  commercially 
important,  though  it  often  accompanies  the  oxidized  ores.  In  the  Lake 
Superior  region  it  is  found  in  wide  lodes  disseminated  through  the  lode- 
matter  0.65  per  cent  to  4  per  cent  of  the  whole,  and  even  when  the  lowest 
grade  mentioned,  by  concentrating  can  be  recovered  at  a  profit.  The  con- 


TREATMENT  OF  COPPER  ORES  353 

centrate  or  "  mineral,"  as  it  is  locally  named,  is  produced  in  different 
grades,  ranging  from  30  to  94  per  cent  copper.  Much  of  the  native  copper 
is  pure;  in  other  instances  it  carries  a  little  arsenic. 

Properties  of  Copper. — Its  melting  point  has  been  established  at 
1083°  C.,  and  its  specific  gravity  at  8.89,  while  its  latent  heat  of  fusion  is 
43.3  calories,  and  its  specific  heat  at  170°  C.  is  0.09244  closely  one-tenth 
that  of  water.  Traces  of  oxygen  are  purposely  left,  even  in  the  highest 
grades  of  copper,  since  otherwise  it  would  be  impossible  to  cast  a  sound 
ingot,  as  the  copper  in  the  refining  process  readily  absorbs  gases  that  are 
expelled  during  solidification.  Even  the  best  of  copper  castings  are  some- 
what porous,  and  thus  low  in  electrical  conductivity.  Castings  of  a  con- 
ductivity of  97  Mathiessen's  standard  have  been  made  by  the  addition  of 
small  amounts  of  boron  in  the  ladle  just  before  casting,  thus  deoxidizing 
the  metal  so  that  its  mechanical  properties  are  excellent,  and  it  can  be 
used  for  making  even  intricate  castings. 

Copper  is  mechanically  improved  by  hot  working:  When  it  is  heated  to 
a  bright  red  and  quenched,  maximum  ductility  is  attained,  while  cold 
working  increases  the  tensile  strength,  but  lowers  the  ductility. 

Solid  copper  can  absorb  arsenic  up  to  a  maximum  of  4  per  cent.  In 
small  quantities  it  increases  its  maximum  stress  without  affecting  the 
ductility.  It  has  been  found  that  the  deoxidation  of  arsenical  copper  by 
addition  of  ferro-silicon  greatly  improves  its  qualities. 

The  physical  properties  of  copper  fall  into  two  classes,  viz.,  electric 
and  mechanical,  and  the  treatment  best  suited  to  attain  the  one  is  undesir- 
able for  the  other.  Pure,  soft  dense  metal  has  the  highest  conductivity, 
but  it  is  weaker  for  use  on  transmission  lines. 

THE  EXTRACTION  OF  COPPER  FROM  ITS  ORES 

Copper  may  be  extracted  from  the  ore  by  dry  or  by  wet  methods. 
By  the  dry  method  the  ore  is  smelted,  the  process  being  one  of  igneous 
fusion.  By  the  wet  or  hydro-metallurgical  methods  the  copper  is  leached 
from  the  ore.  The  striking  point  of  difference  between  the  two  methods  is 
that,  in  the  first,  we  melt  the  entire  ore,  effecting  then  a  separation  of  the 
copper  from  the  worthless  part,  while  in  the  wet  method  we  act  upon  the 
copper  alpne,  leaving  the  greater  part  of  the  ore  in  the  original  condition. 

The  Dry  or  Igneous  or  Pyrometallurgical  Methods. — Probably  more 
than  90  per  cent  of  the  world's  production  of  copper  is  by  smelting.  The 
methods  of  smelting  vary  with  the  nature  of  the  ore.  We  may  divide  them 
into  the  following: 

(1)  The  Smelting  of  Oxidized  Ores  that  may  contain  a  little  copper,  or 
of  concentrates  of  native  copper  in  blast-furnaces  and  in  reverberatory 
furnaces  is  done  for  the  production  of  a  crude  copper  called  blister  copper. 


354  COPPER  ORES  AND  THEIR  TREATMENT 

When  performed  in  the  blast-furnace  the  process  resembles  the  smelting 
of  iron  ore  to  produce  pig  iron.  For  fine  ores  or  concentrate  the  rever- 
beratory  furnace  is  preferred,  since  in  the  blast-furnace  much  flue  dust  is 
produced. 

(2)  The  smelting  of  ores  containing  sulphides  of  copper  and  iron  in  the 
blast-furnace  or  the  reverberatory  is  to  yield  a  copper-and-iron  sulphide, 
called  "  matte."  The  amount  of  matte  is  dependent  on  the  amount  of  sul- 
phur per  cent,  so  that  if  part  of  the  sulphur  is  expelled  by  roasting,  or  by 
being  burned  off  or  volatilized  in  the  blast-furnace,  then  less  matte  is  formed. 
As  much  as  70  to  80  per  cent  of  the  sulphur  may  be  got  rid  of  in  this  way. 
The  copper  is  concentrated  into  a  small  amount  of  matte  as  compared 
with  the  original  bulk  of  the  ore.  This  matte  has  to  be  further  treated 
to  obtain  blister  copper. 

Ordinary  Matte  Smelting. — When  roasted  ores  are  smelted,  this  may  be 
called  ordinary  matte  smelting:  About  10  per  cent  of  the  charge  is  coke, 
added  to  cause  its  melting.  If  the  ore  is  imperfectly  roasted  or  is  smelted 
raw,  then  less  coke  may  be  used,  since  the  sulphur  in  the  ore  burns,  pro- 
ducing heat;  also,  this  burning  of  the  sulphur  is  in  itself  a  kind  of  roasting. 
Thus,  again,  the  amount  of  matte  produced  is  materially  reduced. 

Pyrite  Smelting. — The  smelting  of  raw  ore  in  the  blast-furnace  is  called 
pyrite  smelting  because  there  is  much  iron  or  copper  pyrite  in  the  ore. 

Collectors. — The  blister  copper  produced  in  the  oxidizing  smelting  is 
to  take  up  or  collect  within  itself  any  gold  or  silver  present  in  the  ore; 
the  same  is  true  of  the  matte ;  so  that  we  say  both  blister  copper  and  cop- 
per matte  are  collectors  of  gold  and  silver. 

In  reverberatory  smelting  the  elimination  of  sulphur  is  less,  say  25 
per  cent  of  the  sulphur  in  the  charge ;  so  this  way  of  smelting  is  not  suited 
to  the  treatment  of  raw  ore.  Much  material  for  smelting  has  been  con- 
centrated and  so  is  fine.  It  is  in  good  condition  for  cheap  roasting  in  one 
of  the  mechanical  roasters  already  described. 

Fine  Concentrates. — This  fine  product,  if  treated  in  a  blast-furnace, 
would  produce  much  flue  dust;  so  for  such  material,  the  reverberatory  is 
preferred,  due  to  its  large  capacity,  its  quiet  condition  of  smelting,  and  to 
the  cheaper  fuel  needed. 

The  Wet  or  Hydrometallurgical  Methods. — In  these  the  copper  is 
obtained  from  the  crushed  ore  in  water  solution,  either  with  or  without 
the  aid  of  other  solutions.  The  ore  may  have  first  to  be  roasted.  From 
this  water  solution  the  copper  is  precipitated,  melted  and  refined. 


CHAPTER  XXVIII 
COPPER  BLAST-FURNACE  SMELTING  OF  OXIDIZED  ORES 

This  resembles  the  smelting  of  iron  ores  in  that  metal  is  obtained  in 
metallic  form  in  one  operation.  The  ore,  which  does  not  contain  sulphide, 
is  charged  into  a  blast-furnace  and  smelted  with  coke  for  fuel.  The 
products  are  slag  and  blister-copper,  the  latter  being  metallic  copper  con- 
taining impurities  that  have  been  taken  up  hi  smelting,  much  as  carbon 
and  silicon  are  absorbed  in  iron  smelting. 

Blast-furnace  Plant  for  Oxidized  Ores.— Fig.  190  is  a  plan  and  Fig.  191 
an  elevation  of  a  blast-furnace  building  suited  to  the  smelting  of  oxidized 
copper  ores.  It  has  two  floors  or  levels,  the  upper,  called  the  charge-floor, 
and  the  lower,  the  slag-floor.  The  ground  at  the  right  drops  away  and 
furnishes  a  place  for  a  dump.  The  ores  and  fluxes  are  stored  in  bins  on  the 
ground  at  the  charge-floor  level,  and  are  brought  in  weighed  charges  to 
the  furnace  door,  the  sill  of  which  is  flush  with  the  charge-door.  This  door 
is  shown  also  hi  Fig.  65.  The  slag  and  blister-copper  are  withdrawn  near 
the  bottom.  Behind  the  furnace  is  seen  the  pipe  or  blast-main  by  which 
air  is  conducted  to  the  wind-box  at  the  tuyeres.  The  furnace  stack  extends 
above  the  roof  and  at  the  side  branches  to  a  down-take  leading  to  a  dust- 
flue.  Here  much  of  the  dust,  as  in  the  dust-catcher  of  the  iron  blast- 
furnace is  removed.  This  dust-chamber  terminates,  as  shown  in  thejfront 
of  the  plan  view,  at  a  stack  which  takes  away  the  residual  gases  at  a  high 
level.  On  the  plan  we  also  see  the  boiler  and  engine  which  drive  the  fur- 
nace blower  at  a  pressure  of  12  to  15  oz.  per  square  inch,  equal  to  24  to  30  in. 
of  a  mercury  column.  Waste  slag,  or  sweepings,  carrying  copper,  are 
returned  to  the  feed-floor  by  a  platform  elevator  shown  in  the  corner  of  the 
furnace-room. 

In  Fig.  65  is  a  view  of  the  cupola  blast-furnace  used  for  the  production 
of  copper  from  oxidized  copper  ores. 

The  crucible  contains  the  molten  contents  of  the  furnace,  the  copper 
below  and  tlie  lighter  slag  floating  upon  it.  There  are  two  tap-holes  and 
two  spouts;  the  lower,  close  to  the  bottom,  is  to  remove  the  molten  copper; 
the  upper,  a  few  inches  higher,  is  to  withdraw  the  slag.  From  time  to  time, 
as  slag  or  copper  accumulates,  it  is  withdrawn  by  piercing  a  hole  through 
the  clay-stopping  of  the  tap-holes  by  means  of  a  pointed  steel  tapping-bar. 
The  flow  is  arrested  by  thrusting  into  the  opening  a  plug  of  clay  stuck  on 

355 


356        COPPER  BLAST-FURNACE  SMELTING  OF  OXIDIZED  ORES 


FIG.  190. — Sectional  View  of  Small  Smelting  Plant. 


FIG.  191.— Plan  of  Small  Smelting  Plant. 


COPPER-SMELTING  PLANT  357 

the  end  of  a  button-headed  stopper-rod  or  dolly.  The  slag  is  received  into 
a  fore-hearth  mounted  on  wheels  through  which  flows  the  escaping  furnace 
slag.  This  slag  carries  with  it  drops  of  copper  not  settled  out  in  the  fur- 
nace. The  molten  slag  quite  fills  the  fore-hearth,  crusting  over,  but  main- 
taining a  cavity,  where  the  drops  of  copper  settle  out.  The  slag  overflows 
at  the  spout  at  the  opposite  end  into  a  slag-pot,  Fig.  200,  set  to  receive  it. 
When  the  tap-hole  is  stopped  by  a  plug  of  plastic  clay,  the  slag  flow  ceases, 
and  an  empty  slag  pot  replaces  the  full  one. 

The  copper  is  received  in  a  "  bullion  mold,"  which,  after  filling,  stands 
until  the  copper  has  solidified.  The  ingot  is  then  dumped  out  and  the 
mold  again  used.  The  furnace  shown  is  a  round  one,  36  in.  diameter  inside, 
thus  having  a  bosh  or  enlargement  of  6  in.  on  the  side. 

In  operation,  the  furnace  is  kept  full  to  the  feed-door  with  alternate 
layers  of  fuel  and  charge.  The  blast  rises  through  the  column  of  materials 
(a  distance  of  7  ft.)  and  passes  off  through  the  down-take,  which  has  suffi- 
cient draft  to  take  away  the  gas  and  smoke  and  also  the  air  that  enters 
the  feed-opening  or  door.  The  blast  enters  under  .pressure  causing  an 
intense  combustion  of  the  coke  and  the  fusion  of  the  charge.  The  copper 
reduced  by  the  glowing  coke,  collects  in  drops  and  finds  its  way  to  the 
bottom  of  the  crucible,  while  the  gangue  of  the  ore,  fluxed  by  the  addition 
of  iron  and  limestone,  forms  a  fusible  slag. 

The  smelting  of  the  oxidized  copper  ores,  as  above  described,  was 
formerly  used  in  the  southwestern  United  States  so  long  as  the  oxidized 
ores  lasted;  it  has  been  replaced  by  the  more  efficient  and  cheaper 
methods  of  matte-smelting  either  in  blast-furnaces  or  in  reverberatories. 

The  notable  exceptions  are  those  of  the  copper  country  of  Northern 
Michigan  and  the  large-scale  work  of  the  Union  Miniere  du  Haut  Katanga, 
Southern  Congo. 

LAKE    SUPERIOR    COPPER    COUNTRY    BLAST-FURNACE    SMELTING    OF 

COPPER  SLAG 

Both  tne  cupola  furnace  (Fig.  65)  and  the  rectangular  furnace 
resembling  Fig.  196  are  used.  The  smelting  is  conducted  along  the  lines 
described  for  oxidized  copper  ore  for  the  production  of  an  impure  blister 
copper.  Reverberatory  slags  to  which  has  been  added  native  copper  and 
briquet  ted  material  are  thus  smelted.  The  charge  consists  of  slag  2000  lb., 
limestone  600  lb.,  and  anthracite  coal  (steamer  size  or  about  2J  in.  diameter) 
400  lb.  It  is  the  practice  at  one  works  to  add  much  small  mass  or  native 
copper  to  the  charge,  the  idea  being  that,  as  the  native  copper  in  it  melts 
and  sinks  to  the  crucible  of  the  furnace  in  the  form  of  drops,  it  carries 
down  reduced  copper  with  it.  Elsewhere  the  fine  concentrate  has  been 
thoroughly  incorporated  with  a  quicklime  paste,  briquet  ted  in  a  briquet  ting 


358        COPPER  BLAST-FURNACE  SMELTING  OF  OXIDIZED  ORES 


SMELTING  AT  UNION  MINIERE  DU  HAUT  KATANGA  359 

press  and  the  briquettes  treated  in  cylinders  under  steam  pressure  for 
twenty-four  hours.  This  renders  them  quite  hard,  and  in  an  acceptable 
form  for  smelting. 

The  products  of  the  furnace  are  slag  of  less  than  1  per  cent  copper  and 
"  cupola  blocks,"  an  impure  blister  copper,  which  is  sent  to  a  reverberatory 
furnace  for  refining. 

It  is  to  be  noticed  that  the  low  percentage  of  copper  in  the  slag  is  due 
to  the  use  of  the  anthracite  coal.  This  has  an  intenser  reducing  action 
than  the  coke.  However,  it  slows  down  the  f urnace  and  the  two  fuels 
are  to  be  used  in  judicious  proportions,  because,  as  the  anthracite  is  cut 
down  and  the  coke  increased,  the  furnace  will  run  the  faster.  This  use 
of  coal  suggests  itself  in  cases  where  oxidized  ores  have  elsewhere  to  be 
smelted  in  a  blast-furnace. 

The  cupola  blocks,  referred  to  above,  which  constitute  the  product  of 
the  blast-furnace  smelting  of  the  slag  from  the  reverberatory  refining- 
furnace,  are  melted  and  refined  to  produce  a  low-grade  copper  called 
"  casting-copper."  The  cupola-blocks  are  impure  and  contain  so  much 
arsenic  that  it  is  practically  impossible  to  remove  it  all.  Other  impurities 
(iron  and  sulphur)  are  eliminated. 

SMELTING  TO  BLACK  COPPER  BY  THE  UNION  MINIERE  DU  HAUT  KATANGA 

The  company  smelts  oxidized  copper  ore  from  the  great  superficial 
deposits  at  Lubumbashi,  near  Elisabeth ville,  Belgian  Congo,  producing 
black  or  blister  copper  as  already  described  under  the  heading  "  Blast- 
furnace smelting  of  oxidized  ores."  There  are  six  furnaces,  Fig.  192, 
each  44  in.  wide  and  20  ft.  long,  the  largest  thus  far  made  for  the  treatment 
of  oxidized  copper  ores.  They  have  two  tiers  of  jackets,  the  lower  side 
jackets  being  10  ft.  high  and  2  ft.  wide.  The  upper  ones  of  the  same  width 
are  7  ft.  4  in.  high.  From  the  tuyeres  to  the  feed-floor  is  18  ft.,  and  the 
side-bosh  is  14  in.,  making  the  furnace  shaft  above  the  lower  jackets  72  in. 
wide,  for  good  reduction.  At  the  left-hand  end  is  a  trapped  slag-spout  for 
the  continuous  flow  of  the  slag;  at  the  other  end  is  a  slag-spout  to  be  used 
in  case  the  furnace  stops  and  the  slag  has  to  be  tapped  off.  At  one  side,  and 
nearer  the  right-hand  end,  is  the  bullion  spout  and  tap-hole  for  removing 
the  bullion  (black  copper)  from  the  bottom  of  the  crucible.  To  the  top 
of  the  crucible  from  the  slag-floor  it  is  7  ft.  2  in.,  so  there  is  sufficient  room 
for  the  fore-hearth  beneath  the  trapped  slag-spout.  The  bustle-pipe,  21  in. 
by  42  in.,  in  cross-section  is  rectangular  in  order  to  save  building  space. 
The  furnace  is  surmounted  by  a  closed  top  (not  shown)  as  in  Fig.  196,  with 
feed-doors  the  full  length  of  the  side. 


CHAPTER  XXIX 
BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

MATTE  SMELTING 

If  we  smelt  raw  sulphide  ores  of  copper  and  iron  in  the  blast-furnace 
just  described,  using,  say  10  per  cent  of  fuel,  with  added  fluxes  to  form  a 
fusible  slag,  we  form  an  artificial  sulphide  or  matte  of  23  to  25  per  cent 
sulphur.  The  copper  that  was  in  the  ore  and  the  iron  from  the  charge 
enter  the  matte,  but  the  quantity  of  matte  so  formed  is  but  little  less  than 
that  of  ore  originally  put  into  the  furnace.  If,  however,  we  first  roast  the 
ore,  the  quantity  of  sulphur  present,  and  consequently  the  amount  of 
matte  made,  is  less,  and  the  ratio  of  the  ore  to  the  matte  may  be  five  or 
ten  to  one.  In  the  matte  will  be  the  copper  as  a  sulphide,  and,  in  forming, 
the  matte  will  take  up  the  silver  and  gold  of  the  ore.  By  this  operation 
we  collect  the  precious  metals  in  a  product  one-fifth  to  one-tenth  the  original 
ore,  the  matte  being  termed  a  "  collector."  It  then  can  be  further  treated 
to  convert  it  into  metallic  copper  carrying  the  precious  metals.  By  the 
the  process  of  electrolytic  refining,  the  gold  and  silver  are  eventually 
separated  from  the  copper.  A  charge  suited  to  matte-smelting  methods 
would  therefore  consist  of  roasted  ore  retaining  7  per  cent  sulphur  together 
with  oxidized  copper  and  copper-free  ores  containing  gold  and  silver 
added  to  recover  the  precious-metal  content.  It  would  also  carry  fluxes 
to  make  a  fusible  slag  and  to  supply  iron  (if  needed)  for  the  matte. 

The  products  of  the  furnace  are  slag  and  matte.  The  former  is  the 
result  of  the  union  of  the  silica  in  the  charge  with  the  various  bases,  chiefly 
iron  oxide  and  lime.  The  latter  is  the  complex  artificial  sulphide  pro- 
duced by  the  sulphur  in  the  charge  combining  with  copper  and  iron.  The 
affinity  of  sulphur  for  copper  is  greater  than  for  iron,  and  it  takes  the  former' 
first;  then  if  it  needs  iron  it  takes  that  also,  until  a  compound  of  Doth  has 
been  formed  that  contains  approximately  25  per  cent  sulphur.  Any 
further  iron  present  enters  the  slag  as  ferrous  oxide,  and  as  the  ratio  of  the 
iron  thus  available  to  the  other  bases  varies  so  will  the  slag  vary  in  com- 
position; but  the  principal  requirement  is  that  the  quantity  of  silica  be 
enough  to  form  a  fusible  slag.  Slags  of  25  to  40  per  cent  silica  are  common 
in  copper-matting  practice,  and  the  40  per  cent  limit  is  sometimes  exceeded 
when,  for  economy  in  smelting,  it  is  desired  to  use  as  little  flux  as  possible. 

360 


COPPER  MATTE  SMELTING 


361 


In  the  matte-smelting  operation  the  object  is  to  collect  the  metals, 
copper,  gold,  and  silver,  into  a  small  amount  of  that  complex  artificial 
iron-copper  sulphide  called  matte.  Now,  in  the  pyritic  smelting  of  copper 
ores,  using  an  excess  of  air  and  little  fuel,  much  of  the  sulphur  (say  75  to  80 
per  cent)  is  dissipated  or  volatilized  so  that  a  small  amount  only  is  left  to 
form  matte.  The  same  result  is  attained  by  first  roasting  the  ore,  only  in 
such  case  much  more  coke  must  be  used,  as  compared  with  that  needed  in 
pyritic  smelting.  Blast-furnace  smelting  is  suitable  to  ore  in  lump  form, 
whether  unroasted,  or  as  the  product  of  heap-roasting. 


SECTION   B 


FIG.  193.— Messiter  Bedding  System. 


Ores  for  Matte-smelting. — Ores  suitable  for  matte-smelting  are  the 
oxidized  ones  containing  some  sulphur  and  ores  that  have  been  roasted; 
to  which  may  be  added  silicious  and  oxidized  ores  containing  gold 
and  silver.  Ores  that  require  to  be  first  roasted  are  better  roasted  in  lump 
form  in  heaps  or  stalls  or  sintered,  for  the  reason  that  the  blast-furnace, 
because  of  its  strong  air  currents,  is  not  suited  to  smelting  fine  ore.  The 
amount  of  matte  made  (matte  fall)  depends  upon  the  quantity  of  sulphur 
in  the  charge,  and  to  get  sufficient  concentration  (little  matte  from  much 
ore)  the  sulphur  is  kept  low.  To  the  ore  above  described  is  added  lime- 
stone and  iron  ore  as  fliix,  and  a  quantity  of  fuel  equal  to  10  to  15  per  cent 
of  the  charge. 


362 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


THE  MESSITER  SYSTEM  OF  BEDDING 

Instead  of  storing  ore  in  pockets,  or  bedding  upon  the  ground  as  above 
described,  the  Messiter  system  of  bedding  and  reclaiming  ore  is  coming 
into  increasing  use  for  large  plants.  As  shown  in  Fig.  193,  it  constitutes 
part  of  a  complete  belt-conveying  system.  The  ore,  coarse-crushed  at  the 
sampling  and  crushing  mill,  is  there  separated  by  trommels  into  fine  and 
coarse  ore,  the 'fines  going  to  pockets  for  roasting. 

There  are  three  bedding  floors  side  by  side  upon  which  the  coarse  ore 


FIG.  194. — Robins-Messiter  Reclaiming  Machine. 

is  deposited,  each  bed  when  completed  being  of  triangular  shape  in  cross- 
section,  375  ft.  long,  capable  of  holding  10,000  tons.  By  an  incline  con- 
veying belt  the  ore  is  delivered  upon  the  cross-conveying  belt,  No.  1,  at 
the  left-hand  end  to  the  spreading  belts -Nos.  2,  3,  and  4,  by  me|ns  of  a 
tripper,  each  of  these  three  extending  from  end  to  end  of  the  bed.  A 
tripper  on  each  spreading  belt  travels  back  and  forth  over  the  extent  of  the 
bed  at  the  rate  of  400  ft.  a  minute,  dropping  its  load  in  thin  layers  upon  the 
bedding  floor  beneath.  In  this  way  lot  after  lot  is  distributed  until  the 
bed  is  completed. 

To  take  up  or  reclaim  the  ore,  a  reclaiming  machine  Fig.  194  is  used. 
It  is  a  traveling  frame,  R,  spanning  the  width  of  a  bed,  and  a  trench  con- 
taining a  conveying  belt  that  takes  away  the  ore  delivered  to  it  by  the 


MESSITER  SYSTEM  OF  ORE  BEDDING 


363 


reclaiming  machine,  as  shown  in  Section  A.  Under  the  forward  edge  of 
the  bridge  is  a  scraper-conveyor,  operating  in  a  steel  trough  that  has  a 
flat  bottom  and  a  vertical  back  plate.  The  flights  of  the  conveyor,  sweep- 
ing along  this  trough,  carry  the  ore  to  the  end  of  the  machine  to  drop  it 
upon  the  trench  conveyor-belt.  A  triangular  harrow  covering  the  cross- 


m 


FIG.  195. — Copper-matting  Blast  Furnace. 


section  of  the  bed  and  set  at  a  proper  inclination  has  a  slow  but  powerful 
action  back  and  forth  sufficient  to  dislodge  the  material  which  then  rolls 
down  within  reach  of  the  flights  below.  Both  harrow  and  flight  con- 
veyor are  motor-driven.  Another  motor  advances  the  bridge  into  the 
bed  at  the  desired  rate  or  may  move  it  backward,  when  it  is  transferred 


364 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


to   another  bed.    This  is  effected  by  a  long  transfer  car   (see  Fig.  302) 
moving  in  a  trench  at  one  end. 

THE  COPPER-MATTING  BLAST-FURNACE 

For  producing  matte  from  copper-bearing  ores,  whether  these  are  to 
be  smelted  after  roasting,  or  treated   raw,    by  the    method  of   pyritic 


FIG.  196. — View  of  Copper  Blast-Furnace. 


I 


smelting,  we  use  the  furnace,  Fig.  190,   already  described,  or  one  of  the 
rectangular  type,  Figs.  195  and  196. 

Fig.  195,  at  the  left,  represents  a  transverse  sectional  elevation  of  a 
furnace  of  42  by  120  in.  interior  hearth-dimensions,  having  eighteen 
tuyeres,  nine  at  each  side,  and  a  capacity  of  150  tons  of  charge  daily. 
Fig.  196  is  a  perspective  view  of  a  similar  but  larger  furnace,  differing  from 
Fig.  195  in  having  a  trapped  slag-spout,  more  fully  shown  in  Fig.  196. 


COPPER  BLAST-FURNACE 


365 


The  sole-plate  of  the  furnace  rests  on  jack-screws,  and  can  be  lowered 
and  set  aside  when  it  is  desired  to  make  repairs  or  to  clean  out  the  furnace. 
It  is  protected  from  the  action  of  the  molten  matte  and  slag  by  a  9-in.  lining 
of  firebrick.  In  Fig.  195  are  seen  crucible  plates  which  rest  upon  the  sole- 
plate.  These  are  lined  with  18-in.  of  brick.  The  hollow  jackets  filled  with 
water,  shown  in  Fig.  196,  extend  down  to  the  sole-plate  and  the  water- 
cooling  is  sufficient  protection  from  the  action  of  the  molten  materials. 
The  sole-plate  within  the  furnace,  however,  is  covered  by  the  brick  lining. 
The  jackets,  shown  separately  in  Fig.  197,  are  at  least  9  ft.  high,  and  in  the 
furnace  represented, there  are  two  of  them  on  each  side,  and  one  at  each 
end.  At  one  end  the  jacket  is  shorter,  and  the  space  below  is  filled  with 
a  water-cooled  tap-jacket  through  which  the  slag  is  withdrawn.  In  Fig. 
195  the  longitudinal  view  shows  the  arrangement  of  a  furnace  with  three 


FIG.  197. — Water-jackets  for  Copper- Matting  Blast-furnace. 

jackets  at  each  side,  and  two  jackets  at  each  end.  The  small  jackets  are 
easily  handled  and  replaced.  In  Fig.  197  the  inlets  for  water  are  at  half 
the  height  of  the  bosh,  and  the  water  outlets  are  at  highest  point  to  keep 
them  full  of  water.  They  are  tied  or  clamped  together  with  heavy  angles, 
but  in  the  other  figures  with  I-beams. 

A  water-cooled  trapped  spout  is  used  in  connection  with  the  furnace  as 
indicated  in  Fig.  196.  Through  it  flow  the  slag  and  matte.  Before  the 
slag  can  overflow  it  must  fill  the  spout  and  cover  the  outlet  or  tap-hole 
through  the  jacket.  It  thus  prevents  the  escape  of  the  blast,  and  flows  in 
a  regular  stream  as  fast  as  it  forms  within  the  furnace.  The  furnace  shown 
in  Fig.  195  is  arranged  differently.  There  is  a  spout  and  a  water-cooled 
tap-jacket  at  the  end  of  the  furnace  through  which  the  slag  is  removed, 
while  the  matte,  as  it  accumulates,  is  removed  by  a  spout  and  a  side  tap- 
hole  at  the  level  of  the  crucible-bottom.  In  this  case,  the  separation 


366 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


between  the  slag  and  matte  is  affected  within  the  furnace;  in  the  former 
case,  where  the  trapped  spout  is  used  (since  matte  and  slag  issue  together) 
they  are  separated  outside  the  furnace  in  the  fore-hearth  or  settler,  Fig. 
199  or  195. 

The  transverse  view,  Fig.  201,  shows  the  side-jackets.  These  have 
brackets  or  knees  riveted  to  them  and  rest  on  I-beams  that  are  secured 
to  the  columns.  Thus,  when  the  sole-plate  is  removed,  the  jackets  remain 
in  place.  The  distance  between  the  side- jackets  is  42  in.  at  the  tuyeres, 
and  66  in.  at  the  top.  The  bosh,  or  enlargement,  is  thus  12  in.  on  the  side. 
Above  the  jackets  are  the  cast-iron  distributing  plates,  forming  the  sills  of 
the  feed-doors.  The  feed-doors  in  the  opposite  long  sides  of  the  furnace 
make  it  accessible  from  end  to  end,  not  only  for  feeding  and  trimming 


FIG.  199. — Portable  Forehearth  or  Settler. 

the  charge,  but  for  cutting  with  chisel-bars  the  accretion  or  scaffolding 
that  may  form  on  the  interior  surface  of  the  jackets. 

The  portion  of  the  furnace  above  the  feed-floor  level,  called  the  stack  or 
top,  is  of  brick  supported  by  a  deck-plate  or  mantel-plate  of  I-beams 
resting  on  the  cast-iron  columns  that  extend  down  into  the  foundation. 
The  upper  portion  of  the  stack  is  a  hood  of  sheet-steel  terminating  in  a 
pipe  that  extends  through  the  roof  of  the  furnace  building.  Sometimes  a 
branch  pipe  leads  from  the  hood  to  a  dust-chamber  where  the  dust  is 
collected. 

The  bustle-pipe,  by  which  the  blast  at  a  pressure  of  f  to  2  Ib.  is  brought 
to  the  furnace,  extends  around  three  sides  and  connects  by  the  sheet- 
metal  pipes  to  the  tuyeres.  The  tuyere  is  6  in.  diameter  and  has 
a  6-in.  screw-cap  into  which  is  inserted  a  nipple  with  a  cap  having 


ACCESSORIES  OF  THE  BLAST-FURNACE  367 

a  mica-covered  peep-hole,  through  which  the  condition  of  the  fur- 
nace can  be  observed.  At  the  branch  above  the  tuyere  is  shown  a  slide- 
valve.  In  Fig.  196,  just  below  the  bustle-pipe,  is  a  waste  launder  to  receive 
the  overflow  from  the  jacket,  and  below  it  is  a  3-in.  water-supply  pipe 
branching  to  each  jacket,  and  to  the  water-cooled  trapped  spout  at  the 
front. 

ACCESSORIES  OF  THE  BLAST-FURNACE 

The  fore-hearth,  made  of  cast-iron  plates,  4  by  6  ft.  inside  dimensions, 
is  lined  with  a  layer  of  brick,  and  is  mounted  on  wheels  so  that  it  can  be 
quickly  set  aside  and  a  new  one  put  in  the  place  when  needed.  The 
slag  and  matte  flow  into  it  at  one  end  and  keep  it  full  of  molten  slag.  At 
the  other  end  the  slag  flows  out.  The  matte  settles  on  the  way  and  col- 
lects in  the  bottom  of  the  fore-hearth,  and,  when  accumulated,  is  tapped  at 
the  tap-hole  and  spout  seen  at  the  side.  Meanwhile  the  slag,  flowing  from 


FIG.  200.— Two-wheeled  Slag-pot. 

the  fore-hearth,  is  caught  in  slag-pots,  Fig.  200,  and  taken  to  the  edge  of 
the  dump  and  poured.  The  slag  cools  on  the  surface  of  the  fore-hearth 
and  forms  a  crust  from  beneath  which  the  molten  slag  flows.  Crust  forms 
also  at  the  sides  and  bottom,  and  becomes  gradually  thicker;  and  after 
several  days  becomes  so  thick  that  the  molten  part  of  the  interior  is  too 
small  to  permit  of  a  good  separation  of  the  matte  from  the  slag.  When 
this  results,  the  fore-hearth  is  pried  back  on  the  wheels,  and  replaced  by 
another.  In  the  furnace,  Fig.  196,  at  the  middle  side-jacket,  another 
tap-hole  furnished  with  a  spout  is  seen.  This  is  generally  kept  closed, 
but  is  opened  when  it  is  desired  to  empty  the  hearth  of  the  matte  and  slag. 

The  Slag  Pot,  as  shown  in  Fig.  200,  is  used  for  small  furnaces  both 
in  making  blister  copper  and  for  matting.  It  is  hand-drawn  and  emptied 
at  the  edge  of  the  dump. 

Ladle  Cars. — These  are  used  for  slag  and  matte  in  the  operation  of  a 
large  furnace  and,  as  seen  in  208,  may  be  in  a  train  drawn  by  an  industrial 
locomotive. 


368 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


Fig.  201  is  the  type  of  fore-hearth  used  for  large  furnaces.  It  lasts  in- 
definitely. It  is  lined  on  sides  and  bottom  with  basic  brick  to  resist  the 
corrosive  action  of  the  slag.  The  slag  flows  into  it  at  the  back  from  the 
trapped  furnace  spout  and  escapes  in  a  steady  stream  at  the  slag  spout  in 
front,  to  be  taken  away  in  a  slag  pot  to  the  dump,  see  Fig.  210.  At  the 
side  is  seen  the  matte-tap  whence  from  time  to  time  the  matte  is  tapped 
out  into  a  large  ladle  and  taken  to  the  converter. 


FIG.  201. — Stationary  Fore-hearth. 


FIG.  203.— Positive-pressure  Blower. 


Blowers. — For  furnishing  the  blast  to  the  furnace  the  positive  bfrast 
rotary  blower  is  used,  as  shown  in  Fig.  203,  and  in  section  at  204.  It 
will  be  seen  that,  by  the  rotation  of  the  two  impellers  meshing  into 
one  another  so  that  the  air  cannot  escape  backward  between  them,  the  air 
must  be  delivered  to  the  furnace  in  a  positive  manner  and  not,  as  in  a  fan 
blower,  be  able  to  escape  backward  when  the  pressure  rises  sufficiently. 
The  air  delivered  is  reckoned  at  the  displacement  per  revolution.  If  the 
pressure  is  greatly  increased  there  is  a  backward  leakage  of  air, 


REACTIONS  IN  THE  COPPER  BLAST-FURNACE 


369 


called  the  slip.  Two  boxed-in  cut  gears  outside  the  blower  keep  the 
impellers  exactly  in  mesh.  Blowers  are  often  direct-driven  by  electric 
motor;  but  the  one  described  above  is  belt-driven. 


BLAST-FURNACE   CONDITIONS 

The  diagram,  Fig.  205,  shows  the  reac- 
tions  and  conditions  within  the  shaft  when  a 
furnace  is  smelting  roasted  ore  with  a  full 
amount  of  coke.  It  will  be  understood  that 
the  furnace  is  full  to  the  top  with  solid 
charge  and  the  diagram  shows  the  course  of 
the  descending  charge  and  of  the  rising 
gases. 


350  c. 


1       t       t        I 

Gases,  -  CO2-f  N2 
little  CO,         Inert"  ot 
Neutral  atmosphere 


Descending  Solid  Charge 
preheated  by  ascending 
I  hot  gases  • 

Mi* 


Gases  rapidly  ascending 

CO  2  -H   N2+  CO 
reducing  atmosphere 

,  t  •  M  f 

Focus 
ne  of  Com: 
bust  ion 

Oxidatioi 
of  carbonaceoi 

fuel 

pace  filled  wil 
solid  Incandes-i 
cent  cok( 


feres 


FIG.  204. — Cross-section  of  Blower. 


FIG.  205. — Section  of  Furnace 
in  Pyritic  Smelting. 


LARGE  COPPER-MATTING  BLAST-FURNACES 

The  tendency  of  late  years  has  been  to  increase  the  size  of  copper- 
matting  blast-furnaces.  Increase  in  width  would  require  higher  blast 
pressure  to  drive  the  air  to  the  center  of  the  furnace;  hence  increased  capaci- 
ty had  to  be  gained  by  increasing  the  length  of  the  furnace.  At  the  same 
time,  to  supply  more  air,  the  blast-pressure  has  been  increased  in  some 
cases  to  40  oz.  or  2J  Ib.  per  sq.  in.  A  furnace  56  by  180  in.  under  these 
conditions  smelts  400  tons  of  ore  daily.  For  the  matte  to  properly  settle 
from  the  slag,  with  so  large  a  flow,  a  cylindrical  fore-hearth,  Fig.  201, 
has  been  used,  16  ft.  diameter  by  5  ft.  deep  exterior  dimensions,  lined  with 
hard  firebrick;  for  a  low-grade  matte,  basic  brick  is  sometimes  used.  In 
the  crusted-over  pool  of  molten  slag,  the  separation  is  effected.  The 
molten  contents  of  the  furnace  flow  into  it  at  one  side  and  the  slag  flows 
out  at  the  opposite  side.  From  time  to  time  the  fore-hearth  is  tapped  at 


370 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


FIG.  206. — Longitudinal  Section  of  Copper-matting  Blast-furnace. 


FIG.  207. — Transverse  Section  of  Copper-matting  Furnace. 


OPERATION  OF  THE  COPPER  BLAST-FURNACE  371 

the  lower  tap-hole,  and  5  to  10  tons  of  matte  are  drawn  into  a  ladle,  for 
further  treatment  at  the  converter.  The  lengthening  of  the  furnace  has 
been  carried  so  far  that,  at  the  Washoe  plant,  Anaconda,  Mont.,  a  furnace 
51  ft.  long,  having  1600  tons  daily  capacity,  has  been  for  some  time  in 
operation,  and  recently  one  87  ft.  long  and  3000  tons  daily  capacity  has 
been  built  and  operated.  The  first  furnace  has  two  fore-hearths,  each  16  ft. 
diameter  and  the  second  one  three  of  that  size. 

Figs.  206  and  207  are  elevations  of  a  type  of  copper  blast-furnace  of 
the  new  plant  of  the  Granby  Cons.  Co.  at  Anyox,  B.  C.  It  is  54  in.  wide 
at  the  tuyere  level  and  30  ft.  long.  The  crucible  is  supported  upon  a 
water-cooled  base-plate,  and  its  sloping  bottom  slants  both  ways  to  the 
trapped  spout  at  the  side.  The  jackets  are  in  two  tiers.  The  space 
between  them  widens  to  6  ft.  above  and  at  the  throat  this  is  contracted  to 
4  ft.  6  in.  At  each  side  are  feed-boxes  and  at  the  right  is  seen  a  charge- 
car  dumping  its  load  into  one  of  them.  When  ready  the  charge  is  pushed 
into  the  furnace  by  a  cylinder-operated  plunger  that  forms  the  back  of  the 
box.  It  falls  to  the  top  of  the  charge  several  feet  below.  By  varying  the 
speed  of  the  plunger,  the  cascade  of  ore  can  be  made  to  fall  nearer  or 
farther  from  the  opposite  side,  as  experience  suggests. 

REGULAR  OPERATION  OF  THE  COPPER  BLAST-FURNACE 

Starting  the  Furnace. — Since  a  blast-furnace  is  water-jacketed  the 
operation  of  warming  it  is  a  simple  one.  The  firebrick  lining  of  the  cru- 
cible is  dried  and  warmed  by  several  hours'  heating  with  a  wood  fire.  The 
end  and  side  tap-jackets  are  removed  to  permit  the  air  to  enter  to  the  fuel. 
When  the  hearth  is  hot  the  wood  ashes  are  scraped  out  and  a  fresh  fire 
of  wood,  filling  the  crucible  a  foot  deep,  is  started,  wood  of  uniform  sized 
pieces  for  uniform  burning  being  selected.  Upon  the  wood  is  placed  char- 
coal, and  upon  the  charcoal  coke,  until  the  surface  is  1J  to  2  ft.  above  the 
tuyeres.  The  fire  is  increased  uniformly  and  regulated  by  checking  the 
draft  at  the  front  and  admitting  air  at  the  rear  as  required.  When  the 
coke  is  thoroughly  ignited,  the  furnace  is  ready  for  charging.  The  brick- 
lined  fore-hearth,  Fig.  201,  is  warmed  while  warming  the  furnace.  The 
wood  is  placed  carefully  against  its  walls,  leaving  the  center  clear  for  the 
air  to  reach  the  fuel,  so  that  the  burning  may  proceed  actively.  As  the 
wood  burns,  charcoal  and  ashes  accumulate,  and  are  shoveled  out,  since 
otherwise  they  form  a  layer  through  which  the  heat  does  not  penetrate. 

Suppose  the  charge  of  ore  and  flux  to  be  2000  lb.,  and  that  we  intend,  as 
in  regular  work,  to  use  with  it  12  per  cent  coke  or  240  lb.  per  charge.  We 
put  in  a  layer  of  240  lb.  coke,  then  one  of  500  lb.  slag.  This  is  followed  by 
a  half  dozen  charges  each  of  240  lb.  coke  alternated  with  1000  lb.  slag. 
Next  we  put  in  a  half  dozen  charges  of  240  lb.  fuel  and  2000  lb.  slag,  so 


372  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

that  the  slag  when  melted  shall  entirely  fill  the  fore-hearth.  We  now  begin 
feeding  the  regularly  calculated  charges  and  required  fuel.  At  this  time 
the  blast  is  admitted,  gently  at  first,  and  increasing  during  a  half  hour, 
after  which  the  furnace  should  be  in  full  blast.  Extra  men  should  now 
assist  in  charging  to  rapidly  fill  the  furnace.  At  the  slag-floor,  before  the 
blast  is  turned  on,  the  tap- jackets  and  the  tapped  spout  are  put  in  place, 
and  all  openings  closed  with  a  clay  plugging  mixture.  This  may  be  ob- 
tained from  a  neighboring  bank  if  of  suitable  quality,  or  may  be  made  from 
coarsely  ground  fire  brick  mixed  with  clay. 

As  the  smelting  proceeds,  by  looking  into  the  tuyeres,  we  see  that  the 
slag  is  rising  to  their  level.  We  then  open  the  tap-hole  and  permit  the 
slag  to  flow  into  and  quickly  fill  the  fore-hearth.  The  excess  steadily 
overflows  to  the  slag-pots  set  to  catch  it,  or  it  may  be  granulated  and 
removed  by  water. 

When  feeding  the  furnace,  care  is  taken  to  distribute  the  charge  evenly, 
not  feeding  coarse  ore  in  one  place  and  fine  in  another.  Unless  we  exercise 
care  in  this  regard  we  have  irregular  operation,  blast  and  flame  coming  up 
in  one  place  and  the  charge  looking  dead  in  another.  When  this  begins 
to  occur  we  load  the  active  places  with  charge,  feed  lightly,  and  use  coarser 
material  where  there  is  little  action. 

We  may  adopt  either  the  intermittent  or  the  continuous  method  of 
removing  the  slag  from  the  furnace.  In  the  intermittent  method,  as 
arranged  in  Fig.  195,  the  slag  is  tapped  from  time  to  time  as  it  accumulates 
as  already  described,  taking  care  that  it  does  not  gather  in  such  quantity 
as  to  rise  to  and  run  into  the  tuyeres,  or  to  "  slag  "  them,  as  it  is  called. 
By  the  continuous  method,  the  slag  and  matte  flow  continuously  from  the 
furnace  through  the  trapped  or  open  spout,  see  Fig.  196.  The  molten 
products  enter  a  fore-hearth  and  the  separation  of  slag  from  matte  is  there 
made. 

Referring  to  the  view  of  a  smelting  plant,  Figs.  190  and  191,  the  furnace 
being  filled  to  the  feed-doors,  we  have  a  7-ft.  smelting  column  (distance 
from  tuyeres  to  feed-door) .  As  the  charge  smelts  and  the  molten  materials 
are  withdrawn,  the  surface  gradually  sinks,  making  room  for  further 
additions.  The  coke  is  first  added  in  a  layer  over  the  surface,  and  upon  it  is 
spread  the  weighed  charge.  The  air,  under  pressure  from  the  blowers, 
driven  into  the  furnace  at  a  pressure  not  less  than  J  Ib.  or  12  oz.  per  square 
inch,  burns  the  descending  coke  mostly  at  the  tuyeres.  The  resulting  gas 
with  the  sulphur  dioxide  from  the  burning  sulphur  in  the  charge  appears 
as  a  whitish  smoke  mingled  with  dust  mechanically  carried.  This  passes 
from  the  furnace-top  directly  into  the  air,  or  to  a  dust-chamber  and  thence 
to  the  stack.  The  sulphur  remaining  unites  with  the  copper  and  a  part  of 
the  iron  and  forms  a  matte  or  copper-iron  sulphide.  The  matte,  in  forming 
comes  into  contact  with  the  gold  and  silver  contained  in  the  ore  of  the 


DISPOSAL  OF  SLAG  AND  MATTE  373 

charge  and  absorbs  them.  The  coke  reduces  the  iron  not  needed  for  the 
formation  of  the  matte  to  ferrous  form,  and  the  ferrous  iron,  with  the  lime, 
alumina,  and  other  bases,  combines  with  the  silica  to  form  a  slag,  fluid  at 
the  high  temperature  prevailing  at  the  tuyeres. 

The  molten  slag  and  matte  flow  from  the  furnace  to  the  fore-hearth 
where  the  separation  is  effected,  and  the  supernatant  slag,  freed  from 
matte,  escapes  by  an  overflow  spout,  and  is  received  into  slag-pots  and 
conveyed  to  the  dump,  see  Fig.  208.  Another  means  of  disposing  of 
the  slag  is  to  allow  the  stream  of  slag  from  the  fore-hearth  to  fall  into  a 
launder  and  be  caught  by  a  horizontal  jet  of  water  to  break  it  into  drops 
and  cool  it  in  granules  about  wheat-size.  The  granules  are  carried  away 
by  the  water,  in  a  cast-iron  lined  launder  to  the  dump.  The  matte  is 


FIG.  208.— Pouring  Slag. 

tapped  from  the  fore-hearth  as  it  accumulates,  through  a  tap-hole  near  the 
bottom,  and  flows  over  the  matte  spout  shown  at  the  right  of  the  trans- 
verse section,  Fig.  195. 

COPPER  MATTE 

Matte  is  an  artificial  sulphide  formed  in  smelting  as  a  result  of  the  union 
of  sulphur  with  bases.  Iron  sulphide  (FeS),  such  as  is  used  in  the  making 
of  hydrogen  sulphide  in  the  laboratory,  is  the  simplest  form.  To  produce 
it  in  small  quantities,  a  covered  assay  crucible  may  be  filled  with  shingle- 
nails  and  brought  to  a  white  heat  in  a  wind-furnace  and  roll-sulphur  added 
gradually  until  the  content  fuses.  The  sulphide  is  then  poured,  and  broken 
for  use.  In  smelting  a  charge  containing  sulphur,  scrap-iron  will  take  up 
the  sulphur  and  form  matte.  If  copper  oxide  or  copper  sulphide  is  present 
in  the  charge,  the  sulphur  takes  the  copper  to  form  the  matte  in  preference 
to  taking  the  iron.  When  the  copper  is  exhausted,  the  excess  of  sulphur 
expends  any  combining  power  that  may  remain  by  taking  iron.  We  thus 
have  a  copper-iron  sulphide,  called  copper  matte.  If  lead  or  nickel  are 
present  in  the  charge,  they  partly  enter  the  matte.  Magnetic  iron  oxide, 
taken  from  the  charge,  also  enters  the  matte.  Thus  we  get,  finally,  a  com- 
plex compound,  as  the  following  table  shows : 


374 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


COMPOSITION   OF  COPPER  MATTE 


Cu, 
Per  Cent. 

s, 

Per  Cent. 

Fe, 
Per  Cent, 

Fe04, 
Per  Cent. 

Sp.  Gr. 

Reverberatory  furnace  (Anacenda,  Mont.)  .  . 
Reverberatory  furnace  (Butte,  Mont.)  
Blast-furnace  (Butte  Mont  ) 

60.76 
29.41 
36  15 

23.25 
23.70 
23  38 

11.43 

25.35 
24  97 

1.13 
12.60 
8  51 

5.4 

4.8 
5  1 

Blast-furnace  (Jerome,  Ariz.)  

55.00 

23.96 

13.85 

2.58 

5.3 

Blast-furnace  (Elizabeth,  Vt.)  

21.36 

22.95 

41.03 

10.44 

4.7 

Blast-furnace  (Sudbury,  Canada)  

24.54 

23.24 

28.65 

7.32 

5.1 

Sudbury  matte  contains  also  15.56  per  cent  nickel  replacing  copper. 
It  will  be  noted  that  the  percentage  of  sulphur  (23  to  24)  is  approximately 
the  same  in  all  cases. 

Melting-point  of  Mattes.— 32.6  per  cent,  875°  C.;  49.7  per  cent,  955°; 
61.2  percent,  1070°;  71.1  per  cent  (white metal) ,  1121°,  5  Cu2SFeS;  80.1 
percent  (pimple  metal),  1098°  remaining  after  separating  bottoms;  me- 
tallic copper,  1083°.  Bottoms  contain  Cu,  60  per  cent,  Pb  33  per  cent. 

COPPER-FURNACE  SLAGS 

Variation  in  slag  composition  is  permissible  in  copper-smelting,  the 
requirement  being  that  the  slag  be  fluid  to  flow  from  the  tap-hole  of  the 
furnace.  Slags  having  the  maximum  content  of  silica  and  of  bases,  as 
shown  below,  are  employed  successfully  in  the  blast-furnace. 

In  silver-lead  smelting  practice,  such  variations  are  not  allowable. 
Slags  varying  from  the  composition  found  in  practice  to  be  satisfactory, 
even  though  they  be  fluid  and  run  well,  carry  off  both  lead  and  silver.  In 
copper  practice  such  slags  would  be  clean  and  free  from  copper. 

THE   COMPOSITION  OF  SLAGS 


SiO2, 
Per.Cent. 

FeO, 
Per  Cent. 

AUG., 
Per  Cent. 

CaO, 
Per  Cent. 

MgO, 
Per  Cent. 

ZnO, 
Per  Cent. 

BaO, 
Per  Cent. 

Minimum 

20 

2 

2 

2 

0 

0 

0 

Maximum 

57 

70 

18 

40 

8 

20 

42 

A  slag  low  in  silica  could  not  carry  much  of  the  alkaline-earth^  bases, 
and  would  be  high  in  iron,  and  of  a  specific  gravity  over  3.7.  A  silicious 
slag  would  work  well  with  a  heavy  limy  base,  and  would  have  a  specific 
gravity  of  3.5.  Since  the  separation  of  slag  from  matte  results  from  the 
difference  in  specific  gravity  of  the  two  substances,  we  expect  a  better 
separation  the  lighter  and  more  silicious  the  slags.  Matte  takes  up  zinc 
sulphide,  where  much  is  present,  and  becomes  lighter,  so  in  this  respect, 
zinc  is  detrimental  to  effective  separation. 


CALCULATION  OF  CHARGE 


375 


CALCULATION  OF  CHARGE  FOR  MATTE  SMELTING 

A  low  sulphur  charge  may  consist  of  roasted  ore,  oxidized  copper-ore 
and  silicious  ores  containing  gold  and  silver.  The  requirement  is  that  the 
charge  contain  copper-bearing  ore,  and  enough  sulphur  to  form  with  the 
copper  a  suitable  matte  that  will  take  up  the  gold  and  silver  that  the 
charge  contains.  Enough  flux  is  added  to  the  charge  to  make  a  suitable 
slag,  and  10  to  15  per  cent  coke  or  charcoal  to  smelt  the  mixture. 

The  products  from  the  furnace  are  slag  and  matte,  the  former  being  the 
result  of  the  union  of  the  silica  of  the  charge  and  the  fuel,  with  the  bases 
that  are  present.  A  part  of  the  sulphur  in  the  charge  is  volatilized  by  the 
heat  of  the  furnace,  but  a  large  part,  still  remaining,  combines  with  the 
copper  and  a  part  of  the  iron,  to  form  the  complex  sulphide  called  matte. 
The  iron  not  needed  for  the  matte  enters  the  slag.  Since  the  copper-fur- 
nace slag  may  vary  within  wide  limits,  we  use  a  silicious  ore  where  silica 
is  abundant  and  a  basic  one  where  plenty  of  iron  is  present,  or  in  treating 
basic  ores. 


CHARGE-SHEET.     REGULAR   MATTE    SMELTING 


Name  of  Ore. 

H,0. 

WEIGHT. 

Cu. 

SiOt. 

Fe+Mn. 

CaO+MgO 

S. 

Wet. 

Dry. 

Per 

Cent. 

wt. 

Per 

Cent. 

Wt. 

Per 

Cent. 

Wt. 

Per 

Cent. 

Wt. 

Per 

Cent. 

Wt. 

Roasted  ore..  .  . 
Limestone  
Coke  (10  per  ct) 

3.0 
0 
0 

c 

1030 

C 
?u  and  1 

] 

1000 
300 
130 

Cu  in. 

Ju  in  m 
?e  in  m 

re  in  m 

10.0 

*lag  = 

itte  = 
itte  = 

itte  = 

100 

25.0 
4.2 
7.2 

250 
12 
9 

30.0 
1.2 

300 
"2 

10.0 

atte  = 
itte  = 

100 

52.0 
1.8 

j 

ind  Ft 

156 
2 

100 
4 

271 
For  matte  = 

For  slag  = 

302 
129 

158 
3  in  m 
s  in  mi 

100 
25 

96 
225 

173 
Cu 

75 
3 

129 

225 

Slag. 

In  the  Slag. 

Matte. 

SiO, 

=   35  per  cent. 

FeO  =  173  X® 

=222 

S 

=23  per  cent. 

FeO+CaO 

=   55  per  cent. 

CaO 

=  158 

Cu+Fe 

=  69  per  cent. 

Other  bases 

=    10  per  cent. 

Actual  FeO  +  CaO  =330 
Needed                      =425 

Cu  +Fe     6S 

S0«] 

|    =3  (factor). 

100  per  cent 

_     A  e 

Xv 

CaO  too  little          =   45 
~  =1.57  (factor)  and  271  X1.57  =425  of  FeO+CaO. 


Above  is  given  a  charge  calculation,  in  which  the  problem  is  to  treat  a 
single  roasted  ore,  producing  a  slag  of  predetermined  composition,  and  a 
matte  that  will  take  up  the  copper  that  is  present.  Limestone  is  the  only 
flux  to  be  used.  The  charge  is  of  a  size  to  fill  the  charge-car  or  buggy  in 
which  it  is  brought  to  the  furnace. 

We  will  adopt  1000  Ib.  as  a  weight  of  the  roasted  ore,  having  the  com- 
position Cu  10  per  cent,  8162  25  per  cent,  Fe  30  per  cent,  and  roasted  so 


376  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

that  10  per  cent  sulphur  remains.  This  is  to  be  smelted  with  limestone 
containing  SiO2  4  per  cent  and  CaO  52  per  cent  to  produce  a  slag  of  SiC>2 
35  per  cent  and  bases  (FeO  and  CaO)  55  per  cent,  together  90  per  cent, 
leaving  10  per  cent  to  allow  for  other  elements.  The  slag  has  been  chosen 
of  this  composition  as  one  that  has  been  found  to  work  well.  The  coke 
has  12  per  cent  ash  that  consists  of  Si(>2  60  per  cent,  Fe  10  per  cent,  and  CaO 
15  per  cent.  These  figures,  calculated  to  the  coke,  are  Si02  7.2  per  cent, 
Fe  1.2  per  cent,  and  CaO  1.8  per  cent. 

A  metallurgist,  accustomed  to  types  of  ore,  knows  approximately  how 
much  flux  he  needs.  Suppose  we  decide  upon  300  Ib.  flux.  For  the  calcu- 
lation we  enter  on  the  charge  sheet  the  1000  Ib.  ore,  the  300  Ib.  limestone, 
and  10  per  cent  of  these  or  130  Ib.  coke  in  the  column  of  dry  weight. 
When  the  exact  figures  have  been  computed,  the  wet  weights  may  be 
inserted  in  the  adjoining  column,  using  the  figures  for  per  cent  given  in  the 
column  marked  H^O.  The  percentage  of  ore,  flux,  and  fuel  are  then 
written  in  the  appropriate  columns,  and  the  corresponding  weights,  cal- 
culated to  the  nearest  pound,  are  written  in  and  the  totals  added. 

Beneath,  and  at  the  left  of  the  sheet,  tabulate  the  slag  composition. 
Find  the  ratio  of  base  to  silica,  which  in  this  case  will  be  1.57  to  1.  On 
the  right  of  the  sheet  write  the  matte  composition.  We  know  it  will 
carry  23  per  cent  sulphur,  and  roughly  69  per  cent  copper  and  iron.  Also 
find  the  ratio  of  sulphur  to  base,  which  here  is  1  to  3,  or  the  factor  3. 

Let  us  first  consider  the  sulphur.  Experience  shows  that  in  regular 
matte  smelting  we  can  depend  upon  a  loss  by  volatilization  of  20  to  40 
per  cent  sulphur.  We  take  25  per  cent  as  an  average,  and  thus  75  per  cent 
of  the  sulphur  is  left  to  form  matte.  This  is  75  Ib.,  and  multiplied  by  the 
factor  3  indicates  that  225  Ib.  Cu  and  Fe  together  are  needed  to  satisfy 
the  sulphur. 

The  slag  produced  is  calculated  by  dividing  the  weight  by  the  per  cent 
of  silica,  expressed  decimally,  or  271-7-0.35  =  770  Ib.  Allowing  0.5  per  cent 
copper  for  the  slag  (and  in  good  work  it  should  not  exceed  this,  the  weight 
so  lost  is  4  Ib.,  leaving  96  Ib.  to  enter  the  matte.  Subtracting  this  weight 
of  copper  from  the  total  225  Ib.  of  copper  and  iron  together  needed  for  the 
matte,  we  get  129  Ib.  iron  entering  the  matte,  out  of  the  total  302  Ib.  in 
the  charge.  The  remainder  (173  Ib.)  is  available  for  the  slag.  But  the 
iron  existing  in  the  charge  as  ferric  iron  is  reduced  to  ferrous  form,  and  we 
have,  in  the  ratio  of  atomic  weights,  56  parts  Fe  equal  72  of  FeO,  or  173  Ib. 
Fe  equal  222  Ib.  FeO.  To  this  we  add  the  158  Ib.  CaO,  making  380  Ib. 
of  the  two  bases.  Multiplying  the  silica  (271  Ib.)  by  the  factor  1.57  we 
find  we  need  425  Ib.  of  the  bases  FeO  and  CaO,  so  that  we  have  a  deficit 
of  45  Ib.  Now,  since  the  limestone  consists  approximately  half  of  CaO,  we 
need  to  add  90  Ib.  limestone  to  the  charge,  making  in  all  390  Ib.  as  the 
required  amount.  Erase  where  needed,  and  re-calculate  the  charge. 


PYRITE  MATTE  SMELTING  377 

throughout.  This  time  we  should  come  within  a  few  pounds  of  the  correct 
amount.  As  long  as  it  is  within  10  Ib.  it  is  close  enough,  since  variations 
in  the  ores,  imperfect  weighing,  and  variation  in  the  amount  of  sulphur 
volatilized  easily  exceeds  such  differences.  When,  by  experience,  we  have 
learned  the  actual  percentage  of  volatilization  we  substitute  it  for  that  above 
assumed.  The  actual  percentage  of  copper  and  iron  in  the  matte  is  taken 
in  the  same  way. 

The  grade  of  the  matte  in  copper  is  learned  from  the  ratio  of  the  sul- 
phur (75  Ib.)  to  the  copper  (96  Ib.)  or  23  to  29  per  cent.  In  the  same  way 
we  compute  from  the  respective  weights  the  percentage  of  SiO2,  FeO  and 
CaO,  their  aggregate  being  90  per  cent. 

The  metallurgist  seldom  can  count  on  the  slag  and  matte  coming  from 
the  furnace  precisely  as  calculated.  There  is  a  little  variation  due  to  the 
causes  already  mentioned.  When  the  slag  from  a  newly  calculated  charge 
comes  down,  a  sample  should  be  taken  and  a  rapid  determination  made  for 
Cu,  SiO2,  FeO,  and  CaO.  As  an  approximate  rule,  to  increase  an  ingre- 
dient of  the  charge  a  given  percentage,  add  to  it  the  fractional  part  ex- 
pressed by  its  ratio  to  the  remainder  of  the  100  per  cent.  Thus,  if  analysis 
gives  33  per  cent  and  we  wish  to  increase  it  to  35  per  cent,  then  to  the  -f$  of 
271  =  16.4  Ib.  add  |f  (i)  of  16.4  Ib.,  making  the  total  silica  to  be  added 
25  Ib. 

PYRITE  MATTE  SMELTING 

This  consists  in  treating  in  a  blast-furnace,  such  as  that  shown  in  Fig. 
196,  sulphide  ore  consisting  largely  of  pyrite  and  chalcopyrite.  The  ore 
carries  gold  and  silver,  which  are  recovered  in  the  copper-bearing  matte 
produced.  No  preliminary  roasting  is  given  the  ore,  and  the  smelting  is 
conducted  in  such  a  way  that  70  to  80  per  cent  of  the  sulphur  is  burned  in  the 
furnace  while  the  remainder,  uniting  with  iron  and  the  copper,  forms  the 
matte  which  acts  as  collector  for  the  gold  and  silver.  A  slag  is  formed 
from  the  silica  of  the  gangue  and  the  bases  of  the  ore  and  flux.  At  times 
when  the  quantity  of  base,  especially  iron,  is  large  it  is  necessary,  in  order 
to  make  a  suitable  slag,  to  add  silicious  ore.  The  matte  and  slag  flowing 
together  from  the  furnace  separate  in  the  fore-hearth. 

It  will  be  noticed  that  the  slow  and  expensive  preliminary  roasting  of 
the  sulphide  ores  is  obviated,  and  that  the  amount  of  fuel  needed  is  small 
(1.5  to  6  per  cent)  because  of  the  heat  developed  by  the  burning  of  the 
sulphide.  Pyrite  or  chalcopyrite  contains  iron  that  is  available  both  for 
matte  and  slag,  and  when  the  matte  can  spare  it  for  the  slag  the  iron 
serves  to  flux  the  silica  of  iron-free  ores  on  the  charge.  Iron  ore  or  lime- 
stone acts  in  the  same  way,  and  either  of  them,  though  generally  the 
latter,  may  be  added  for  the  purpose. . 

An  iron  matte  alone  does  not  entirely  collect  the  gold  and  silver  from 


378  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

the  ore-charge,  and  it  has  been  found  that  copper,  to  the  extent  of  0.5 
per  cent  or  more,  should  be  present  to  insure  the  collection  of  these  metals 
in  the  matte.  The  slag  then  will  be  nearly  free  from  the  precious  metals. 
Copper,  therefore,  acts  as  an  efficient  collector. 

As  a  result  of  burning  70  to  80  per  cent  of  the  sulphur  of  the  charge,  there 
remains  only  30  to  20  per  cent  to  form  matte.  The  remaining  sulphur  first 
takes  up  copper,  for  which  it  has  a  greater  affinity  than  for  iron.  It  is  the 
burning  off  of  the  large  amount  of  sulphur  that  enables  one  to  dispense 
with  roasting  and  to  diminish  the  amount  of  matte  produced.  The  matte 
produced  per  ton  of  ore,  or  the  matte-fall,  may  be  expressed  as  a  percentage, 
or  as  a  concentration  of  so  many  tons  into  one  of  matte.  Thus,  with  a 
production  of  200  Ib.  matte  per  ton  of  ore,  we  have  a  10  per  cent  matte-fall, 
or  a  concentration  of  10  into  1.  It  is  desirable  to  concentrate  the  ore  into  a 
small  bulk  of  matte.  To  show  how  much  concentration  is  effected,  both 
in  regular  matte  smelting  and  in  pyrite  smelting,  we  enter  upon  the  fol- 
lowing considerations: 

In  regular  smelting  (with  a  charge  containing  8  per  cent  sulphur,  the 
volatilization-loss  being  25  per  cent,  and  the  matte  to  contain  25  per  cent 
sulphur),  we  have  from  100  Ib.  ore  75  per  cent  of  8  per  cent  =  6  Ib.  sulphur 
to  form  matte.  This  makes  24  Ib.  matte  and  results  in  a  concentration  of 
4.2  into  1. 

In  pyrite  smelting  with  a  charge  containing  30  per  cent  sulphur,  the 
volatilization  loss  being  80  per  cent  and  the  matte  still  to  contain  25  per 
cent  sulphur,  we  have  from  100  Ib.  of  ore  20  per  cent  of  30  per  cent  =  6  Ib. 
of  sulphur  to  form  matte.  This  makes  24  Ib.  of  matte,  the  same  concen- 
tration as  in  the  regular  matte  smelting  just  specified. 

It  will  be  noted  that  the  percentage  of  volatilization,  or  the  amount 
of  sulphur  burned,  varies  with  the  charge.  It  is  low  when  only  roasted  or 
oxidized  ores  are  used,  and  high  for  raw  or  unroasted  ores,  especially  those 
containing  pyrite.  A  defect  inherent  in  pyrite  smelting  is  the  difficulty  of 
regulating  this  loss.  When  the  furnace,  which  is  burning  the  right  amount 
of  sulphur,  begins  to  run  slow  from  any  cause,  the  volatilization  may 
increase  to  the  point  of  burning  the  entire  content  of  sulphur,  so  that  no 
matte  is  produced.  On  the  other  hand,  when  the  furnace  begins  to  run 
fast,  much  matte,  low  in  copper,  is  produced.  The  principal  difficulty  in 
pyrite  smelting  is  the  regulation  of  the  matte-fall. 

REACTIONS  IN  PYRITE  MATTE  SMELTING 

Fig.  209  is  a  cross-section  of  a  matting  blast  furnace  smelting  a  pyritic 
charge,  and  consisting  of  sulphide  ores  with  the  addition  of  enough  quartz 
or  silicious  ore  to  make  a  suitable  slag.  With  this  charge  3  per  cent  of 
coke  is  used,  but  this  does  not  interfere  with  the  reactions. 


REACTION  IN  PYRITE  SMELTING 


379 


At  the  surface  of  the  charge,  which  is  maintained  at  12  ft.  above  the 
tuyeres,  where  the  temperature  may  be  at  250°  C.,  the  heat  drives  off  from 
the  FeS2  a  portion  of  its  sulphur,  leaving  as  Fe3S4.  The  escaping  sulphur 
fume,  encountering  the  air  entering  the 
feed-door,  burns  with  the  characteristic 
blue  flame  to  862. 

By  the  time  the  charge  has  gone 
down  5  ft.  in  the  furnace  where  the 
temperature  is  much  higher,  more  sul- 
phur has  been  expelled,  leaving  FeS. 

At  7  ft.  down,  and  at  a  tempera- 
ture of  925  to  950°  C.,  dissociation 
continuing,  we  find  Fe5S4=4FeS+Fe, 
or  a  condition  in  which  four  equivalents 
of  FeS  hold  one  of  Fe  in  solution,  so 
that,  if  a  sample  of  the  compound  in  a 
molten  condition  could  be  withdrawn 
from  the  furnace  we  would  find  iron 
separating  from  the  iron  sulphide  on 
cooling. 

The  FesS4  at  this  zone  begins  to 
melt  and  falls,  entering  the  silicious, 
porous  structure  or  "  nucleus  "  shown 
above  and  below  the  tuyeres  in  Fig. 
209. 

Within    this    porous     structure    the         FIG.  209. — Section  of  Furnace  in 
downward  trickling  sulphide  is  encoun-  ^tic  Smelting, 

tering  the  rising  air  according  to  the  reaction 

(1)       4FeS     +     Fe     +     13O     =     5FeO     +     4SO2 

4X23,800  5X66,400      4X71,000   =   520,800 

This  is  equivalent  to  1874  pound-calories  per  pound  of  iron  present. 

The  forming  FeO  at  once  unites  with  the  silica  present  to  form  a  molten 
slag  which  continues  its  journey  to  the  crucible.  The  rising  gases,  intensely 
heated  as  the  result  of  the  reactions,  expand,  keeping  the  nucleus  porous. 
The  ore-column,  therefore,  instead  of  resting  on  a  bed  of  burning  coke  in 
the  crucible  as  in  regular  matte  smelting,  is  sustained  upon  a  network  of 
quartz  pieces  constituting  the  nucleus,  and  which  have  thus  far  escaped 
slagging.  The  nucleus  extends  from  the  crucible  upward  to  the  zone  of 
fusion  of  the  iron  sulphide  5  ft.  above  the  tuyeres,  having  its  greatest 
development  at  or  about  the  tuyeres.  At  the  sides  of  the  furnaces  the  net- 
work is  moving  slowly  downward,  and  is  sustaining  the  portion  which  is 
descending  regularly. 


380  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

Generally  not  all  the  FeS  is  oxidized,  but  a  part,  together  with  any  cop- 
per sulphide  present,  forms  matte  and  in  molten  condition  seeks  the 
crucible. 

In  a  regularly  working  furnace  producing  its  equivalent  slag,  this  slag 
is,  within  limits,  not  changed  in  composition  by  a  change  in  the  amount  of 
silica.  It  must  be  understood  that  the  addition  of  silica  has  the  effect  of 
increasing  the  degree  of  concentration,  that  is,  of  raising  the  grade  of  the 
matte.  There  is  then  more  iron  oxidized  and  slagged  off,  but  the  effect 
is  not  to  make  the  slag  more  irony,  but  simply  to  increase  its  quantity.  If 
too  much  silicious  ore  is  added,  the  excess  remains  undigested  and  chokes 
the  furnace.  If  too  little  is  added,  the  amount  of  matte  increases  while 
its  tenor  in  copper  becomes  less,  and  at  the  same  time  the  amount  of  slag 
decreases,  but  without  altering  its  composition.  Considered  in  another 
way,  the  pyrite  furnace  chooses  its  own  slag. 

As  compared  with  other  kinds  of  smelting,  an  abundance  of  air  should 
be  supplied  to  the  furnace,  not  less  than  300  cu.  ft.  to  1  Ib.  of  sulphur 
present  in  the  charge. 

At  Mt.  Lyell,  Tasmania,  where  great  success  has  been  attained  in 
pyritic  smelting,  two  parts  of  heavy  sulphide,  ore  containing  2  to  2.25  per 
cent  Cu  are  used  to  one  of  silicious  ore  of  70  per  cent  SiC>2  with  a  con- 
centration of  18  to  20  parts  into  one,  producing  a  matte  containing  40 
per  cent  Cu.  Of  the  large  amount  of  iron  present  95  per  cent  is  burnt  or 
oxidized,  the  small  remainder  going  into  the  matte.  Under  given  condi- 
tions, a  rise  from  95  to  96  per  cent  of  iron  oxidized  results  in  an  increase  in 
the  grade  of  the  matte  to  50  per  cent  Cu. 

The  porous  condition  of  the  nucleus  is  practically  preserved  as  the  result 
of  the  reactions  there  taking  place,  while  it  is  the  duty  of  the  metallurgist 
to  see  that  such  loose  and  porous  condition  is  suitably  maintained  else- 
where in  the  shaft  of  the  furnace,  which  must  be  kept  properly  open  both 
below  and  above  this  fiery  net-work.  The  proper  maintenance  of  this 
condition  is  one  of  the  principal  secrets  of  success  in  order  to  avoid  freeze- 
ups  and  to  use  the  minimum  amount  of  coke. 

Pyrite  Smelting  in  Two  Stages. — For  low-grade  ores,  carrying  2  per 
cent  copper,  for  example,  a  concentration  of  ten  into  one  gives  matte  of 
20  per  cent  copper.  By  roasting  the  matte,  or  by  smelting  it  pyritically, 
it  is  possible  to  increase  the  grade  to  40  per  cent  or  more,  and  thisiproduct 
can  be  treated  in  the  copper-converter  and  brought  to  the  grade  of  blister- 
copper.  An  ore  containing  5  per  cent  copper  can  be  smelted  to  give  a 
matte  of  40  per  cent  copper,  so  that  the  second  smelting  with  the  additional 
expense  can  be  omitted.  Now  while  it  would  not  pay  to  smelt  a  copper 
ore  of  a  grade  as  low  as  2  per  cent  copper,  for  the  copper  alone,  if  the  ore 
contained  gold  and  silver  the  recovery  of  these  metals  would  justify  the 
expense  of  smelting. 


CALCULATION  OF  CHARGE 


381 


Two-stage  Smelting  at  Ducktown,  Tenn. — At  the  works  of  the  Tennes- 
see Copper  Co.,  the  process  is  a  two-stage  one,  the  ore  being  smelted  to  give 
matte  of  10  per  cent  Cu  ("  ore-smelting  "),  this  matte  being  re-treated  in 
another  furnace  to  produce  a  35  to  45  per  cent  matte  ("  matte-smelting  "). 
The  slag  from  the  second  furnace,  not  yet  sufficiently  clean  to  throw  away, 
is  remelted  in  the  first  furnace.  The  second  matte  is  then  converted  to 
blister-copper  of  98  per  cent. 

The  furnaces  (of  the  type  shown  in  Fig.  196)  are  196  in.  long  by  56  in. 
wide  at  the  tuyere  level,  and  have  large  circular  fore-hearths,  16  ft.  out- 
side diameter  by  5  ft.  deep,  lined  with  chromite  brick  to  resist  the  corrosive 
action  of  the  low-grade  matte  produced.  The  first,  or  ore-furnace,  treats 
400  tons  of  charge  daily,  with  a  coke-consumption  of  2.1  per  cent.  The 
second  or  concentration  furnace  smelts  280  to  300  tons  of  matte,  using 
3.5  per  cent  coke.  The  following  charge-sheet  gives  details  regarding  the 
charge  and  the  matte  produced  in  both  the  stages,  and  shows  how  such 
charges  are  computed  in  pyritic  smelting. 

CALCULATION  OF  CHARGE  IN  PYRITE  SMELTING 

First  Stage,  the  Ore  Charge. — In  these  calculations  the  quantity  of  coke 
is  so  small  that  no  computation  is  required  for  the  ash.  The  quantity  of 
base  in  the  ore  is  so  large,  and  the  silica  is  so  low  that  it  has  been  necessary 
to  add  silicious  material  (in  this  case  quartz-rock)  to  the  charge  in  order  to 
obtain  a  slag  of  35  per  cent  silica.  The  problem  is  to  compute  the  amount 
of  quartz  to  be  added  to  give  such  a  slag. 

CHARGE    SHEET    I.     ORE-SMELTING    FURNACE. 


Name  of  Ore. 

WEIGHT. 

H,O. 

Cu. 

SiOi. 

Fe+Mn. 

CaO+MgO 

S. 

Wet. 

Dry. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Polk  Co  
Burra-Burra  .  .  . 
Quartz 

Coke  

100 

1000 
3000 
700 

'  'cii 
Cuin 

2.4 
2.1 

n'si'ag 
matte 

24 
63 

20.7 
9.4 
97.0 

207 

282 
679 

34.2 
38.0 

342 
1140 

9.2 
8.3 

Volati 
In 

II 

92 
249 

20.4 
30.3 

890 
43 

204 
909 

87 
=    6 

1168 

1482 

341 
lized  = 

shiir  = 

[atte  = 

1113 
933 
180 

=  81 

Slag 

SiOj=35.0  per  cent. 
FeO+CaO  =55.0  per  cent. 

S=1.3  per  cent. 
Cu  =0.2  per  cent. 
Slag  =  1168-^0.  35  =3300 


Matte 

Fe  +Cu  =65  per  cent. 
S  =25  per  cent. 

°=i*-»6 


Factor 

oo 


.  636 


We  represent  on  the  charge  sheet  the  ores  that  are  to  be  run,  using 
amounts  in  accord  with  the  rate  at  which  the  respective  ores  are  supplied 


382  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

(1000  lb.  Polk  county  ore  and  3000  Ib.  Burra-Burra  ore).  Experience 
shows  that  for  such  a  charge  and  for  the  quantity  of  ferrous  iron  and  lime 
present,  we  may  enter  the  quantity  of  silicious  material  as  700  lb.  We 
use  a  slag  of  35  per  cent  Si()2  and  55  per  cent  FeO  and  CaO,  making  in  all 
90  per  cent.  With  this  slag,  experience  shows  we  may  figure  on  1.3  per 
cent  sulphur  and  0.2  per  cent  copper  for  this  low-grade  matte.  A  little 
zinc,  when  that  element  is  present  in  the  charge,  also  enters  the  slag. 
This  is  shown  to  be  0.3  per  cent. 

The  percentage  of  ingredients  of  the  charge  is  written  and  carried  out 
in  the  respective  columns,  and  the  columns  are  added. 

Beginning  with  the  sulphur,  of  the  1113  lb.  present,  we  have: 

Lb. 

Sulphur  volatilized  (80  per  cent  of  the  total) 890 

Sulphur  in  the  slag  (1.3  per  cent  of  3300  lb.) 43 

Sulphur  left  for  the  matte 180 

1113 

The  matte  is  assumed  to  contain  25  per  cent  sulphur,  65  per  cent  copper 
and  iron,  and  1.7  per  cent  zinc.  We  estimate  that  80  per  cent  of  the  zinc 
will  be  volatilized.  It  is  understood  that  in  smelting  other  ores  than  these, 
the  actual  quantities  of  the  different  elements  in  the  matte  and  slag  will  be 
determined  and  those  figures  substituted  for  the  ones  above. 

Of  the  copper,  0.2  per  cent  of  3300  lb.  or  6  lb.  goes  into  the  slag,  leaving 
81  lb.  for  the  matte.  Multiplying  the  sulphur  for  the  matte,  180  lb., 
by  the  factor  2.6  we  get  the  total  Fe  and  Cu  needed  for  the  matte,  468  lb. ; 
subtracting  Cu  for  matte,  81  lb. ;  leaves  Fe  for  matte,  387  lb. 

But  the  total  iron  in  the  charge  is  1482  lb.,  so  that  we  have: 

Lb. 

Fe  in  matte 387 

Fe  left  for  slag 1095 


Total 1482 

The  iron  in  the  slag  occurs  as  FeO,  so  that  we  must  take  1408  lb.  FeO 
equivalent  to  the  1095  lb.  Fe.  Adding  to  this  the  CaO,  341  lb.,  we  get 
FeO + CaO  =  1749  lb.,  which  multiplied  by  the  factor  0.636,  gives  SiO2  = 
11121b. 

Lb. 

Actual  silica  in  charge 1168 

Silica  needed 1112 

Silica  in  excess 56 

By  erasing  the  trial  amount  700  lb.  of  quartz,  and  substituting  650  lb. 
then  recalculating  the  charge,  we  get  an  approximation  within  10  to  20  lb., 
which  is  accurate  enough  for  practical  purposes.  The  percentage  of  copper 


CONCENTRATION  OF  MATTE 


383 


in  the  matte  is  computed  according  to  the  proportion  159  :  81  ::  25  per 
cent  :  10.4  per  cent. 

Second  Stage,  Matte  Concentration. — This  is  run  with  slag  from  the 
converting  operation,  and  quartz  ore  in  sufficient  quantity  to  produce  a 
slag  of  the  same  composition  as  that  of  the  ore-charge,  except  that  it  has 
1  per  cent  of  sulphur  and  0.7  per  cent  copper. 

CHARGE  SHEET  II.  MATTE-CONCENTRATION  FURNACE 


Name  of  Ore. 

WEIGHTS. 

H«0. 

Cu. 

SiOj. 

Fe+Mn. 

CaO+MnO 

S. 

Wet. 

Dry. 

Per 
Cent. 

Lb. 

Per 

Cent. 

Lb. 

Per 

Cent. 

Lb. 

Per 

Cent. 

Lb. 

Per 
Cent. 

Lb. 

Matte 

3000 
1400 
200 
400 

Cu  in 
Cu  in  n 

10.0 
2.0 

slag  = 
latte  = 

300 
'    8 

97\0 
4.0 
30.0 

1358 
8 
120 

55.0 
55!6 

1650 
220 

53.0 
1.0 

Volat 
IE 

106 
4 

25.0 
1.0 

"  530 

•     43 

750 
4 

Quartz 

Limestone  

Convert  slag.  .  . 
Coke 

ioo 

308 
30 
273 

1486 

1870 

110 
lized  = 

slag  = 

754 

573 

181 

Slag 

SiOs  =35.0  per  cent. 
FeO  H-CaO  =55.0  per  cent. 
S=1.0per  cent. 
Cu=  0.7  per  cent. 
Sign  =1486  -=-0.35=  4300  Ib. 

Factor  |?  =0.636. 


Matte 

Fe  +Cu  =65  per  cent. 

S  =25  per  cent. 

Cu+Fe 


S 


The  charge  is  estimated  as  in  "  Charge  Sheet  I,"  with  a  volatilization 
of  70  per  cent  of  the  sulphur,  as  experience  has  shown  the  result  com- 
monly to  be.  In  the  matte  we  have  : 

Lb.  Per  Cent. 
S  ......................................................    181=25.0 

Cu  .........................  .  ..........................  278  =  83.4 

(181X2.6)  -278  =  Fe  ............................  .   193=26.6 


652=90.0 
Proceeding  with  the  calculation  for  the  quartz  we  have : 

Lb. 

Total  iron. .  1870 

Iron  in  matte 193 

Iron  for  slag 1677 

or  FeO  =  2157,  and  the  total  base  is  2266  Ib. 

This  gives  the  silica  needed,  2266X0.636  =  1509  Ib.  But  we  have 
already  1486  Ib.,  and  the  difference  may  be  made  up  by  increasing  the 
quartz  20  Ib. 

Concentration  of  the  Matte. — In  an  example  above,  we  obtained  a 
matte  10  per  cent  in  copper,  and  with  copper  low  in  the  charge,  the  per- 


384 


BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 


centage  may  be  even  less.  The  matte  is  of  too  low  grade  to  ship  away,  or 
to  bring  to  the  grade  of  blister-copper  in  a  converter.  It  must  be  concen- 
trated to  one  of  higher  grade.  If  we  were  to  heap-roast  the  matte  and 
then  smelt  it  with  silicious  ore  in  a  blast-furnace,  we  should  obtain  a 
small  quantity  of  matte  of  a  high  grade. 

There  is,  however,  the  expense  and  delay  of  the  roasting  to  consider, 
and  it  has  been  sought  to  smelt  the  matte  raw,  with  silicious  ore,  with  the 
idea  of  burning  off  the  sulphur  in  the  blast-furnace.  In  regular  matte- 
smelting,  were  this  attempted,  the  matte  would  run  through  little  dimin- 
ished in  quantity  and  little  changed  in  grade,  but  by  the  new  method, 
using  little  fuel,  an  abundant  blast,  and  silicious  slag,  the  concentration 
can  be  obtained. 

DISPOSAL   OF  THE  SLAG 

The  slag  from  a  blast-furnace,  being  a  waste  material,  is  disposed  of 
in  the  cheapest  way  possible.  In  the  case  of  small  furnaces,  as  it  flows 


FIG.  210. — Electric  Trolley  System  (removing  slag-pot). 


V 


from  the  fore-hearth,  it  is  caught  in  wheeled  slag-pots  (slag-carts),  Fig. 
200,  that  are  taken  to  the  edge  of  the  slag-dump  when  filled  and  poured. 
As  the  dump  grows  the  expense  increases,  and  large  slag  cars,  Fig.  210, 
are  used.  The  cars  are  moved  either  by  horses,  or  an  industrial  loco- 
motive, or  by  trolley. 

Another  cheap  and  favorite  way  is  to  granulate  the  slag.     To  do  this  a 
cast-iron  launder  is  arranged  to  receive  the  slag  as  it  falls  from  the  spout 


BLAST-FLJRNACE  VS.  REVERBERATORY  SMELTING  385 

of  the  fore-hearth.  The  launder  has  a  grade  of  1  in.  to  the  foot,  and  through 
it  water  is  made  to  flow  constantly.  In  addition,  a  horizontal  flattened  jet 
of  water  strikes  the  falling  slag,  instantly  cooling  and  breaking  it  into  gran- 
ules of  various  sizes  averaging  y^  in.  in  diameter.  The  flow  of  water 
carries  the  slag  to  the  dump. 

BLAST-FURNACE  VS.  REVERBERATORY  SMELTING 

Predictions  have  recently  been  made  that  the  reverberatory  was  bound 
to  supplant  the  blast-furnace,  because  of  the  advantage  the  former  pos- 
sessed in  the  treatment  of  the  finer  ore  and  flotation  concentrates.  It  is 
pointed  out,  however,  that  where  ore  is  coarse,  and  where  it  is  possible  to 
avoid  roasting,  the  blast-furnace  has  its  advantages  even  for  sulphide  ores. 

"  There  is  no  doubt  that  at  the  moment,  in  favored  localities  where 
pulverized  coal  or  fuel-oil  can  be  obtained  at  a  much  cheaper  rate  than  coke, 
the  reverberatory  has  the  better  of  the  argument,  but  there  is  always 
something  turning  up  in  favor  of  the  other  side  in  every  controversy.  It 
resembles  the  perpetual  fight  between  armament  and  projectiles  that  we 
are  all  much  more  familiar  with  at  present  than  we  were  prior  to  1914. 

"Both  styles  of  furnaces  have  their  field,  but  for  the  moment,  owing  to 
the  great  increase  in  tonnage  treated  by  oil-flotation  and  the  improve- 
ments in  reverberatory  practice,  the  reverberatory  seems  to  be  gaining 
materially  in  tonnage  treated. 

"The  latest  improvement  in  blast-furnace  practice  is  the  introduction  of 
pulverized  coal  at  the  tuyeres,  the  notable  examples  of  this  being  the 
Tennessee  Copper  Co.  smelter  and  the  International  Nickel-  Co.  's  plant  at 
Copper  Cliff,  Ont.  If  the  experiments  now  being  tried  at  various  plants, 
besides  the  two  mentioned,  prove  that  pulverized  coal  can  be  used  econom- 
ically in  the  blast  furnace,  either  with  or  without  a  proportion  of  coke,  the 
probabilities  being  that  a  certain  proportion  of  coke  will  be  necessary,  the 
existing  interest  in  blast-furnace  practice  will  receive  an  impetus. 

"The  new  plant  being  erected  in  Chile  for  the  Braden  Copper  Co. 
contemplates  the  use  of  the  blast-furnaces  entirely,  using  nodulizing-fur- 
naces  to  prepare  the  charge.  To-day  it  is  about  a  stand-off  in  cost  between 
nodulizing  or  sintering  for  blast-furnace  practice  and  roasting  for  rever- 
beratory practice."* 

Methods  of  Reverberatory  Smelting. — The  essentials  for  successful 
reverberatory  practice  are  self-fluxing  ores,  cheap  fuel,  and  cheap  silica 
brick,  whereas  for  blast-furnace  practice  the  charge  may  be  more  refractory, 
but  must  be  either  sintered,  nodulized  or  come  naturally  in  lumps,  and 
in  addition  the  fuel  and  power  must  be  comparatively  cheap. 

While  the  blast-furnace  in  general  is  the  cheapest  means  of  smelting 
copper-bearing  ores  in  coarse  or  lump  form,  one  objection  to  it  is  that  the 

*  E.  P.  Mathewson. 


386  BLAST-FURNACE  SMELTING  OF  SULPHIDE  ORES 

blast  may  carry  away  5  to  10  per  cent  of  the  fine  dusty  ore.  This  may 
be  settled  as  flue-dust  in  flue  or  dust  chambers,  made  into  briquettes  and 
resmelted,  but  the  additional  expense  should  be  avoided  if  possible.  Ore 
or  concentrate  in  fine  condition  is  better  treated  in  the  quieter  atmosphere 
of  the  reverberatory  furnace.  If  raw  ore  is  treated  in  such  a  furnace  about 
25  per  cent  of  the  sulphur  only  would  be  expelled;  hence,  before  smelting, 
sulphide  ore  would  be  roasted. 


CHAPTER  XXX 


REVERBERATORY    SMELTING 

Two  methods  have  been  evolved  for  the  reverberatory  smelting  of 
copper  ores,  viz.,  the  Welsh  process  and  the  reverberatory  matte  smelting 
process.     The  Welsh  process  possesses  the  advantage  that  it  can  be  used 
on  a  great  variety  of  ores,  the 
final  product  being  a  blister 
copper.       In    reverberatory 
matte  smelting  a  great  ton- 
nage of  roasted  ore  is  put 
through,   and   the   resulting 
product,    which    is,    in    the 
form  of  copper  matte,  must 
undergo  a  further  treatment 
in    the    converter    to   bring 
it    to   the   stage    of    blister 
copper. 


THE  WELSH  PROCESS  OF  RE- 
VERBERATORY SMELTING 

This  consists  in  treating 
copper  ore  (sulphide  and 
oxide  as  well  as  silicious  ore) 
by  a  series  of  roastings  and 
fusions  to  raise  the  grade  of 
the  copper  in  the  product 
finally  to  blister  copper, 
which  is  subsequently  refined 
electrolytically,  as  any  blister 
copper  containing  gold  and 
silver  naturally  would  be. 
The  process  has  the  advan- 
tage that  a  variety  of  ores, 
both  coarse  and  fine,  can  be 
treated  in  a  few  small  furnaces  with  a  small  investment  of  plant. 

Fig.  211  is  a  furnace  lined  with  refractory  material  having  a  shallow 
basin-shaped  hearth.    Square  cross-bars  in  the  fire  at  a  sustain  the  grate 

387 


Dinas  Brick 
Ordinary  Brick 
Flintshire  Brick 
Brasque 
Hearth 


FIG.  211. — Reverberatory  Smelting  Furnace. 


388  REVERBERATORY  SMELTING 

bars  of  1J  in.  square  iron  (not  shown).  A  deep  fire  is  maintained  to  the 
top  of  the  fire  door,  e,  and  some  cinder  is  allowed  to  accumulate  on  the 
grate,  but  so  as  to  keep  the  fire  open.  Any  grate  bar  can  be  pushed  aside 
and  the  cinder  dropped  into  the  ash  pit  b  when  cleaning  the  fire.  Next  to 
the  firebox  is  a  "  bridge  "  2  ft.  6  in.  wide,  to  confine  the  fuel  to  the  firebox. 
A  brick  arch  beneath  it  sustains  the  hearth  foundation  of  brasque  and  on 
this  the  refractory  brick  of  the  hearth.  The  charge  is  thrown  in  at  the  side 
door,  which  is  then  quickly  bricked  up.  The  furnace  is  stirred  from  the 
front  door,  and  the  slag  and  matte  removed  by  rabble.  The  products  of 
combustion  escape  by  a  port  or  opening  in  the  roof  and  by  a  flue  are  carried 
to  the  stack  or  chimney.  The  part  of  the  roof  toward  the  outlet  port  is 
called  the  "  verb." 

SMELTING  OPERATIONS  BY  THE  WELSH  PROCESS 

We  may  divide  the  smelting  operation  into  five  parts: 

(1)  "  Calcining  "  the  Ore. — Sulphide  ore  containing  5  to  15  per  cent 
copper  is  roasted  in  a  hand-reverberatory  roaster  (see  Fig.  29)  until  not 
more  than  5  per  cent  sulphur  is  left. 

(2)  Fusion  of  Ore. — The  roasted  ore  is  charged  in  a  reverberatory 
furnace  with  such  oxidized  copper  ore  as  is  available,  and  melted.     The 
sulphur  contained  in  the  roasted  ore,  with  the  copper  and  some  of  the  iron, 
forms  a  matte  of  35  per  cent  copper,  called  "  coarse  metal."     The  silica, 
uniting  with  the  ferrous  oxide  not  taken  by  the  matte,  and  also  with  the 
alumina,  and  the  alkaline-earth  bases,  forms  a  slag,  fusible  at  the  high 
temperature  of  the  furnace.     The  molten  bath  boils  from  the  escape  of 
trioxide  resulting  from  the  reaction  of  the  ferric  iron  upon  unroasted  fer- 
rous sulphide,  or  from  the  decomposition  of  barium,  lead,  or  zinc  sulphate 
by  silica,  thus: 

(2)  BaS04 + SiO2  =  BaSiO3 + SO3 . 

(3)  PbSO4+SiO2  =  PbSiO3+SO3. 

(4)  ZnSO4 + SiO2  =  ZnSiO3 + SO3 . 

The  sulphuric  anhydride  escapes  as  a  gas,  and  upon  meeting  the 
moisture  of  the  air  at  the  top  of  the  stack,  changes  to  a  white  fume  of  ^2864 . 

(3)  Calcining  Coarse  Metal. — The  coarse  metal  or  matte  that  is  run 
from  the  reverberatory  furnace  in  operation  (2)  into  sand  beds,  is  crushed 
to  pass  a  5-mesh  screen  and  fed  to  another  hand-roaster.     Rich  sulphide 
of  20  to  70  per  cent  copper  also  is  crushed  and  added  to  the  charge.     The 
whole  is  roasted  until  it  contains  not  more  than  5  per  cent  sulphur. 

(4)  Second   Reverberatory   Fusion. — The   roasted   material,   now   of 
35  to  50  per  cent  copper,  is  charged  into  a  fusion-furnace  with  oxidized 
ores  containing  20  to  70  per  cent  copper.     When  the  charge  is  melted 


THE  WELSH  PROCESS  389 

there  results  a  matte  of  75  per  cent  copper,  called  "  white  metal,"  com- 
posed chiefly  of  copper  sulphide.  As  before,  the  silica  contained  in  the 
ore  added  to  the  charge  unites  with  the  ferrous  iron  and  other  bases  to 
form  slag.  This  slag,  however,  having  been  made  from  such  rich  material, 
contains  much  copper  and  is  not  to  be  thrown  away  but  returned  to  another 
charge  in  the  fusion-furnace  of  operation  (2). 

(5)  "  Roasting  "  and  Formation  of  Blister-copper. — The  white  metal  is 
charged  in  large  pieces,  as  broken  when  removing  from  the  sand  molds, 
into  a  reverberatory  fusion-furnace  where  it  is  piled  in  an  open  fashion, 
particularly  near  the  bridge.  It  is  fired  gradually  for  several  hours  with 
an  oxidizing  flame.  A  supply  of  air  is  admitted  at  a  number  of  ports  or 
openings  2.5  in.  square  in  the  roof  over  the  fire-bridge,  and  at  the  sides 
of  the  furnace  near  the  bridge.  The  operation  is  called  "  roasting."  The 
oxidizing  flame,  acting  at  the  surface  of  the  lumps  and  upon  the  drops 
trickling  down,  converts  a  portion  into  cuprous  oxide,  so  that  we  have 
present  copper  both  as  sulphide  and  as  oxide.  Finally  the  heat  is  raised 
and  the  whole  charge  is  melted  down,  according  to  the  reaction: 

(5)          2Cu2O     +     Cu2S     =     6Cu     +     SO2 

2X42,000          20,200  71,000     =-33,200. 

Copious  fumes  of  SO2  issue  from  the  boiling  surface  of  the  molten  charge. 
Slag  rich  in  copper  is  produced,  getting  the  silica  partly  from  the  interior 
walls  of  the  furnace,  partly  from  silicious  but  entirely  oxidized  ore  that  has 
been  added  to  supply  silica.  '  The  slag  is  returned  to  operation  (4) .  Finally 
the  blister-copper  is  tapped.  The  metal  obtains  this  name  from  the  fact 
that,  upon  cooling,  occluded  gas  seeking  to  escape  from  the  molten  metal, 
forms  blisters  on  the  surface  of  the  pigs  of  metal. 

The  blister-copper  now  contains  98  per  cent  copper,  but  also  impurities 
that  must  be  removed  to  make  it  suitable  for  market.  The  refining 
process  is  described  elsewhere.  In  case  the  copper  contains  gold  and  silver, 
taken  from  the  ores  that  supplied  the  copper,  it  is  customary  to  remelt  it, 
pole  it  to  remove  copper  oxide,  and  to  cast  it  into  anodes  for  electrolytic 
refining. 

Treatments  of  the  Bottoms. — Impure  bottoms  are  remelted  and  cast 
into  anodes  for  electrolytic  refining,  in  which  the  impurities  and  gold  are 
separated  from  the  copper;  or  they  may  be  formed  into  an  inferior  grade  of 
copper  (casting-copper)  as  follows:  A  charge  is  put  into  the  blister-furnace 
as  in  operation  (5),  consisting  of  14,000  Ib.  of  75  per  cent  roasted  white 
metal,  21,000  Ib.  raw  white  metal,  8000  Ib.  bottoms  and  1000  Ib.  silicious 
ore.  This  is  melted  down,  and  then  is  added  6000  Ib.  more  roasted  white 
metal.  The  charge  aggregates  50,000  Ib.  This  is  treated  precisely  like 
the  regular  blister-charge,  but  it  yields  a  higher  percentage  of  copper. 


390  REVERBERATORY  SMELTING 


THE  DIRECT  PROCESS  OF  REVERBERATORY  SMELTING 

This  is  a  modification  of  the  Welsh  method  of  producing  blister- 
copper.  Instead  of  "  roasting  "  the  matte  or  white  metal  in  lump  form 
in  the  blister- furnace  or  process  (5),  a  portion  is  ground  to  5-mesh  size  and 
calcined  or  roasted  in  a  separate  roasting-furnace,  of  either  the  hand  or 
the  mechanical  type. 

Into  a  melting-furnace,  called  the  "  blister-furnace/'  is  charged  14,000 
Ib.  of  the  roasted  white  metal,  still  retaining  4  to  6  per  cent  sulphur,  3500 
Ib.  raw  unroasted  white  metal,  4000  to  8000  Ib.  slag  from  a  former  charge, 
and  600  Ib.  silicious  ore  to  unite  with  the  FeO  and  other  base  present. 
When  this  has  been  melted,  6000  Ib.  more  roasted  matte  is  added,  making 
a  total  of  23,500  Ib.  matte  charged.  When  all  is  fused,  the  reaction  begins. 
The  surface  of  the  charge  is  seen  to  be  seething  and  boiling,  and  escaping 
bubbles  of  gas  are  set  free  according  to  the  following  reaction : 

(6)  2Cu2O+Cu2S  =  6Cu+SO2. 

The  bath  is  then  skimmed  to  remove  the  slag,  which  contains  copper 
oxide,  reserved  in  part  for  the  next  charge  and  in  part  sent  back  to  stage 
(4)  of  the  Welsh  process.  From  the  charge  here  specified  there  are  pro- 
duced 75  pigs  weighing  230  Ib.  each,  or  17,250  Ib.  blister-copper,  and  also 
23  pots  of  slag  weighing  400  Ib.  or  9200  Ib.  containing  12  to  15  per  cent 
copper.  Slag  is  removed  several  times  during  the  period.  The  copper, 
when  free  from  slag,  is  tapped  into  a  refining-furnace  at  a  lower  level. 
The  refining  is  done  in  a  14-  by  22-ft.  furnace  carrying  a  deep  charge  of 
copper.  This  charge  is  made  from  copper  scrap,  high-grade  "  mineral  " 
of  over  80  per  cent  copper,  and  mass-copper.  It  is  melted,  and  a  blister- 
copper,  comparatively  free  from  impurities,  is  obtained. 

LARGE-SCALE  REVERBERATORY  MATTE  SMELTING 

This  and  blast-furnace  smelting  are  practically  the  two  methods  of 
treating  copper  ores  in  the  United  States.  Such  ores,  principally  sul- 
phides, are  roasted  and  smelted  in  large  reverberatory  furnaces  for  the  pro- 
duction of  matte  and  of  a  slag  low  in  copper,  which  is  sent  to  waste.  *Fuel, 
rapidly  burned,  maintains  the  furnaces  at  a  temperature  above  the  smelting 
point  of  the  forming  slag.  We  may  divide  them  according  to  the  method 
of  heating  into  (1)  direct-fired  furnaces  having  a  firebox  for  coal  burning; 
(2)  furnaces  fired  with  pulverized  coal,  and  (3)  oil-fired  furnaces.  Except 
at  the  fire-end  they  are  essentially  the  same. 


COAL-FIRED  REVERBERATORY  FURNACE 


391 


(1)     THE  DIRECT  COAL-FIRED  FURNACE 

We  give  at  Fig.  212,  in  plan  and  elevation,  a  large  furnace  showing,  at 
the  "  back  end,"  the  ash  pit  and  the  large  grate,  8  by  16  ft.  area,  and  the 
solid  fire-bridge  immediately  in  front  of  it.  At  the  outlet  end  the  flue 
comes  out  at  the  roof.  There  is  ample  room  in  front  for  skimming  and 
tapping  the  slag.  This  runs  into  a  water-filled  bosh  where  it  is  granulated 
and  swept  away  by  a  powerful  horizontal  jet  of  water.  At  two  points  on 
the  side  the  matte  is  tapped  off  at  the  hearth  level  and  runs  by  gravity 
along  a  matte  launder  to  a  matte  ladle  set  at  a  low  level  to  receive  it.  At 
the  left  are  shown  the  waste-heat  Sterling  boilers  of  300  H.P.  each.  Thence 
the  gases,  having  given  up  much  of  their  heat  pass  on  by  an  under- 


Hearth  116  x  19 
Hearth  Area  2120  sq.  feet 
Grate  Area  110.7  sq.  feet 
Ratio  Grate  to  Hearth  1:19 


FIG.  212. — Coal-fired  Reverberatory  Furnace. 

ground  flue  to  the  main  stack.  In  case  it  is  desired  to  cut  out  the 
boilers  for  repair  the  damper  to  the  branch  "  underground  flue  "  is  opened 
and  the  boiler  damper  is  closed.  Two  boilers  are  heated  by  the  waste  gas 
from  the  furnace,  and  develop,  together,  600  H.P.  The  furnace  treats,  on 
an  average,  275  tons  hi  twenty-four  hours,  producing  a  40  per  cent  matte 
concentrating  4  into  1. 

The  charge  consists  of  hot  roasted  ore  ("  calcines  ")  from  MacDougall 
roasters,  Fig.  77,  and  by  analysis  is  shown  to  be  composed  as  follows: 
Cu  9  per  cent,  FeO  24.4  per  cent,  CaO  2.9  per  cent,  S  8  per  cent,  SiO2 
26  per  cent.  Every  eighty  minutes  a  charge  of  15  tons  is  dropped  into 
the  furnace  near  the  fire-bridge.  This  falls  upon  the  bath  of  molten 
matte  and  slag  that  the  furnace  contains.  It  spreads  in  all  directions, 
and  much  of  it  floats  gradually  toward  the  front.  It  readily  melts  by 
contact  with  the  molten  slag  and  matte  below  and  the  flame  above.  In 


392  REVERBERATORY  SMELTING 

this  great  reservoir  of  heat  there  is  but  little  variation  in  temperature,  and 
the  flame  is  transparent. 

Every  four  hours  45  to  50  tons  of  slag  is  removed  in  fifteen  minutes 
from  the  furnace  and  allowed  to  flow  from  the  front  door  in  a  thick  stream. 
It  is  granulated  by  a  strong  horizontal  stream  of  water  as  it  falls  into  the 
waste-launder.  The  water  sweeps  it  away  to  the  dump  several  hundred 
yards  from  the  furnace.  The  matte  is  kept  at  a  nearly  uniform  level, 
10  tons  being  tapped  out  at  a  time,  while  the  total  amount  in  the  furnace 
is  100  to  150  tons. 

The  action  of  the  slag  upon  the  furnace  is  to  erode  or  scour  it,  but 
because  of  the  width,  the  sides  are  less  acted  upon  than  would  be  the  case 
in  a  narrow  furnace.  To  repair  the  furnace,  both  slag  and  matte  are  drawn 
off  completely,  then  the  side-doors  are  opened,  and  sand  is  thrown  across 
the  furnace  against  the  sides  where  eaten  away  by  the  slag.  They  thus 
are  protected  against  the  inroads  of  the  slag.  A  furnace  runs  six  or  eight 
months  and  then  has  to  be  shut  down  for  the  thorough  repair  of  the  roof, 
walls,  and  bridge.  These  parts  are  of  silica  brick,  the  walls  being  30  in. 
thick,  the  roof  15  in.  Silica  bricks  are  practically  infusible  but  expand 
on  heating,  so  that  allowance  is  made  by  leaving  transverse  slits  in  the  roof. 
These  close  when  the  furnace  is  at  full  heat.  An  important  point  in  effi- 
cient working  is  to  have  the  outlet-flue,  or  "  neck,"  of  the  proper  size.  It 
must  be  large  to  insure  good  draft,  and  yet  retain  the  flame  in  the  furnace. 
In  the  furnace  shown  it  is  60  by  38  in.  or  16  sq.  ft.  area. 

The  temperature  at  fire-bridge  is  1550°  C.;  at  the  flue  end  126o°  C. 
The  slag  leaves  the  furnace  at  1120°  C.  and  drops  to  1060°  at  the  over- 
flow spout  of  the  settler.  Gases  reach  boiler  at  950°  and  leave  at  330°  C. 

Heat  Balance  of 

above  Furnace. 

Per  Cent 

Heat  in  slag 16.2 

Heat  in  matte 3.2 

Heat  lost  in  radiation 11.6 

Heat  lost  in  cooling-bridge  plate 0.2 

Sensible  heat  in  grate  droppings 0.8 

Heat  in  steam  generated  at  boilers 32 . 8 

Heat  in  gases,  passing  the  boilers 13.2 

78.  ot 

The  remaining  22  per  cent  must  be  due  to  the  excess  of  air  over  the  theo- 
retical amount  needed  for  combustion. 

(2)     FURNACES  FIRED  BY  PULVERIZED  COAL 

In  Fig.  213  is  a  sectional  plan  and  elevation  of  a  pulverized  coal-fired 
furnace,  with  a  hearth  116  ft.  long  by  19  ft.  9  in.  wide  and  with  side  walls 
22J  in.  thick.  At  the  firing  end  are  two-  tall  charge  hoppers,  but  the  most 


POWDERED-COAL  FIRING 


393 


of  the  charging  is  done  by  side  hoppers  having  6-in.  feed  pipes  extending 
through  the  roof  near  the  side  walls.  By  opening  a  slide  in  the  feed  pipe, 
calcines  can  be  charged  against  the  wall  as  fast  as  the  material  melts  down. 
At  the  firing  end  is  an  18-in.  blast  pipe  from  two  blowers  with  five  air  jets 


wr.j 

LONGITUDINAL  SECTION 

FIG.  213. — Reverberatory  Furnace  Fired  with  Pulverized  Coal. 


• 27 

TYPICAL  SECTION 


Continuous  Feed  Bins 
10'Veed  Pipes  2t"c.  to  C 


^Position  of    ;      A  <t>  A 
Charge  in  |  'L_  t  t  j_V_  _/ 
Furnace 


Boiler 


ELEVATION  OF  BURNER 

MB 


SKIM   AND  TAP  END 


Skimmiug  Block 


Slag  Line 


FIG.  214. — Method  of  Charging  the  Furnace. 

^ 

to  the  furnace.  From  the  pulverized  coal  bin  come  down  five  pipes  sup- 
plying the  coal  by  a  screw  feed  to  the  tops  of  the  air  branches  or  burners. 
The  coal  as  it  drops  into  the  burner  is  atomized  and  blown  into  the  furnace, 
instantaneously  taking  fire  and  filling  the  furnace  with  flame.  A  slag 
launder  at  one  side  of  the  front  end  takes  the  flow  from  the  tap-hole,  while. 


394 


REVERBERATORY  SMELTING 


OIL-FIRED  FURNACES 


395 


a  matte-launder  towards  the  firing  end  takes  away  the  matte  at  a  tap-hole 
set,  as  shown  in  the  sectional  elevation  at  the  hearth  level.  On  the  oppo- 
site side  of  the  furnace  are  spare  matte  launders  to  be  used  in  case  of 
need.  It  wil  be  seen  that  the  furnace  bottom  is  of  sand  fused  in  layers 
by  heavy  firing. 

How  the  furnace  is  charged  is  well  shown  in  Fig.  214.  The  calcine, 
hot  from  the  roasters,  is  banked  along  the  wall  as  at  B  and  as  shown  in  the 
typical  section  of  the  same  figure.  There  is  also  a  transverse  section  show- 
big  the  position  of  the  burners.  It  will  be  seen  from  the  temperatures  given 
in  Fig.  213,  plan,  that  the  banking  is  near  the  hottest  part  of  the  furnace. 


SECT.  ELEV.   OF 


j 

SECTIONAL  ELEVATION  ON  CENTER  LINE 
FIG.  216. — Low-pressure  Oil-burners. 

(3)     OIL-FIRED  FURNACES 

Where  oil  is  the  cheapest  fuel  it  is  to  be  preferred,  and  in  Fig.  216 
are  shown  in  sectional  plan  and  in  longitudinal  and  transverse  sections 
an  oil-fired  reverberatory  furnace.  In  this  case  there  are  six  charge 
hoppers  placed  at  the  zone  of  greatest  heat,  but  no  side  charging.  The 
charge  as  there  melted  down  flows  toward  the  front  end,  the  matte  settling 
out  in  quiet.  Ore  is  brought  to  these  hoppers  by  a  charge  car  from  the 
roasters.  The  side  walls  are  thick  and  sloped  up  above  the  slag  line  at  the 


396  REVERBERATORY  SMELTING 

side.  This  furnace  has  side  doors  to  give  access  to  the  hearth  and  walls 
for  repairs.  Tapping  of  slag  is  done  at  the  front  under  the  outlet  flue, 
and  matte  just  at  the  point  where  the  furnace  begins  to  narrow.  The 
roof  is  high  toward  the  firing  end,  sloping  downward  gradually  to  the 
verb  at  the  front. 

There  are  four  oil  burners  located  well  above  the  slag  line  and  made  as 
shown  in  Fig.  216.  Oil  under  high  pressure  is  brought  by  a  J-in.  pipe  to  a 
burner  tip  or  nozzle,  where  it  is  caught  up  and  atomized  by  the  air  blast, 
and  thus  made  ready  for  instant  burning  with  an  intense  flame. 

Burning  Temperature  of  Oil  in  a  Reverberatory  Furnace. — A  fuel 
oil  of  the  composition  C  85.0  per  cent,  H  12.4  per  cent  and  3.4  per  cent, 
when  completely  burned,  will  yield  10,720  calories  per  pound.  The  prod- 
ucts of  combustion,  including  0.5  Ib.  steam  for  atomizing  the  oil,  will 
be  (see  page  80)  3.11  Ib.  carbon  dioxide;  9X12.4  per  cent— 0.5  H^O  as 
water,  a  total  of  1.618  Ib.  and  in  the  steam  1.44  Ib.,  a  total  of  3.70  Ib. 
oxygen  that  comes  from  13.91  Ib.  of  air  containing  10.71  Ib.  of  nitrogen. 
Hence  we  have 

3. 11  Ib.  carbon  dioxide @  0.364  =  1 . 132 

1 . 618  Ib.  water  vapor ; @  0 . 77    =  1 . 245 

10.71    Ib.  nitrogen .    @  0.281  =3. 009 


5.386 

The  temperature  of  combustion  is  therefore       '       =2000°  C.   nearly. 

5,o  80 

In  practice  some  excess  of  air  must  be  used  so  the  temperature  is  not 
attained.  Thus,  with  the  excess  of  23  per  cent  of  air,  the  theoretical 
temperature  may  be  given  at  1800°  C.;  and  considering  cooling  influences 
not  more  than  1500°  to  1600°  C.  We  may  assume  that  the  gases  leave  the 
furnace  at  1200°  and  the  waste  heat  boilers  at  400°  C.  Due  to  leakage 
of  air  into  the  furnace  and  to  the  cooling  influence  of  the  gases  escaping 
from  the  charge  the  excess  gases  are  increased  to,  say,  50  per  cent.  Sum- 
marizing we  have 

Per  Cent. 

Calorific  value  of  the  fuel 100 . 0 

Distribution  of  heat: 

Taken  away  by  the  matte  and  slag 13.5 

Radiation  from  the  furnace 10 . 0 

23.5 

Taken  up  by  the  boilers 34 . 5 

Radiation  at  the  boilers 7.0 

41.5 
Sent  to  the  stack. .  .35.0 


100.00 


OPERATION  OF  REVERBERATORY  FURNACE  397 


OPERATION  OF  A  LARGE  REVERBERATORY  FURNACE 

These,  whether  coal  fired,  pulverized-coal  fired  or  oil  fired  are  charged 
much  in  the  same  way.  Part  of  the  charge  is  dropped  into  the  furnace 
from  time  to  time  from  hoppers  shown  near  the  firing  end  of  the  furnace ; 
the  rest  is  added  through  charge  tubes  so  as  to  maintain  a  bank  of  ore 
against  the  furnace  side  wall  at  the  hottest  point,  thus  protecting  the 
wall  against  the  intense  heat  and  the  corrosion  of  molten  slag.  Ore 
is  brought  to  the  hopper  as  well  as  to  the  side  wall  hoppers  by  calcine 
cars  placed  on  transverse  overhead  tracks.  However,  the  Cananea 
oil-fired  furnace  is  not  so  fed.  It  is  provided  with  side  doors  through 
which  ore  can  be  thrown  for  fettling  the  walls  to  the  opposite  side  of  the 
furnace. 

Increase  of  furnace  output  depends  upon  the  increase  of  fuel  burned 
with  its  increased  and  intenser  consequent  heat.  It  has  been  found  that 
by  enlarging  the  outlet  flue  to  70  sq.  ft.  as  compared  with  half  that  dimen- 
sion and  by  increasing  the  volume  of  low-pressure  atomizing  air  with  the 
increased  fuel,  the  flame  was  shorter,  began  closer  to  the  burner  nozzle, 
and  was  intenser,  so  that  melting  proceeded  more  rapidly  and  a  tonnage 
formerly  of  400  was  increased  to  700  tons  and  over.  In  one  instance  in  an 
oil-burning  furnace  of  100  by  21  ft.  hearth  area  800  tons  was  smelted 
daily,  using  0.622  bbl.  of  oil  (183  Ib.)  per  ton  of  charge.  Even  this 
figure  has  been  improved  by  charging  the  calcines  promptly  from  the 
roasters. 

In  older  practice  the  custom  was  to  remove  the  slag  intermittently, 
using  a  rabble  to  assist  the  flow.  More  or  less  crust  and  floating  material 
imperfectly  smelted  was  thus  withdrawn.  In  present  practice  where  400 
tons  and  over  is  smelted  per  furnace,  the  breast  has  been  closed  and 
the  tapping  holes  made  of  such  aperture  that  the  flow  is  continuous  and 
the  crust  near  it  is  undisturbed.  In  this  way  the  copper  in  the  waste  slag 
was  cut  from  0.4  per  cent  to  0.3  per  cent  per  ton  of  slag. 

Ribs  are  now  introduced  in  the  roof,  arching  from  the  buck-staves 
between  the  ribs,  thus  prolonging  the  life  of  the  furnace  roof. 

REACTIONS  AND  CALCULATION  OF  THE  CHARGE 

The  sulphur  of  the  charge  unites  with  the  copper  and  iron  until  its  needs 
are  satisfied.  The  matte  thus  formed,  separating  in  drops,  absorbs  the 
precious  metals  contained  in  the  ore,  and  by  the  greater  specific  gravity 
than  that  of  the  slag,  penetrates  downward  to  the  hearth.  The  silica 
of  the  gangue  unites  with  bases,  such  as  FeO,  CaO  and  A12O3,  and  forms  a 
fusible  slag  that  floats  as  a  separate  layer  upon  the  matte. 

When  ore  is  roasted  a  part  of  the  iron  is  oxidized  to  the  ferric  state, 


398  REVERBERATORY  SMELTING 

and  when  the  charge  is  fused  the  ferric  iron  acts  upon  the  unroasted  ferrous 
sulphide  according  to  the  following  equation : 

(7)  FeS+3Fe203+7SiO2  =  7FeSiO3+SO2. 

When  ferric  iron  is  present  in  the  ore,  sulphur  is  eliminated  of  ten.  to  the 
extent  of  25  to  33  per  cent,  and  we  find  less  matte  than  would  be  present 
if  this  reaction  did  not  take  place.  The  iron,  thus  reduced  to  ferrous  form, 
enters  the  slag. 

CHARGE   CALCULATION  FOR  A  REVERBERATORY  FURNACE 

Charge  Calculation. — The  charge  of  a  reverberatory  furnace  generally 
consists  of  hot  calcines  from  the  roasting-furnaces  to  which  limestone 
has  been  added  at  the  time  of  roasting. 

The  charge  for  the  reverberatory  may  be  calculated  as  in  the  example 
under  "  Regular  Matte  Smelting."  Below  is  given  an  example  where 
the  composition  of  the  charge  is  known  and  we  desire  to  compute  the  conse- 
quent composition  of  the  slag  and  matte.  It  is  assumed  that  not  more  than 
30  per  cent  of  the  sulphur  in  the  roasted  ore  will  be  volatilized,  that  approx- 
imately 0.4  per  cent  copper  and  1  per  cent  sulphur  go  into  the  slag,  which 
also  carries  90  per  cent  of  the  three  elements  SiO2,  FeO,  and  CaO;  that 
the  matte  is  to  contain  90  per  cent  of  the  three  elements,  sulphur,  iron, 
and  lime.  The  ore  has  been  roasted  with  10  per  cent  by  weight  of  crushed 
limestone.  The  charge-sheet  is  arranged  as  in  the  table,  and  the  per- 
centages computed  and  added.  The  weight  of  the  slag  of  35  per  cent  SiO2 
would  be  1200  lb.,  containing  1  per  cent  sulphur,  or  12  lb.,  and  this  added 
to  the  30  per  cent  or  42  lb.  of  the  sulphur  burned  off,  will  make  54  lb., 
leaving  86  lb.  to  enter  the  matte  whose  weight  would  be  four  times  this  or 
344  lb.  Some  65  per  cent  of  the  matte  will  be  copper  and  iron.  After 
allowing  for  0.4  per  cent  or  5  lb.  as  being  lost  in  the  slag,  there  remains 
115  lb.  of  copper,  and  108  lb.  of  iron  is  consequently  needed  for  the  matte. 
Deducting  this  from  the  total  iron  there  is  left  294  lb.,  equal  to  378  lb.  of 
FeO.  Let  us  now  assemble  the  items  for  the  slag.  They  are  silica  408 
lb.,  FeO  378  lb.,  and  lime  140  lb.,  constituting  90  per  cent  of  the  slag.  The 
resultant  percentages  are  then  computed.  In  the  same  way  in  theimatte, 
since  the  three  constituents,  sulphur,  iron,  and  copper,  make  90  per  cent, 
we  can  distribute  the  resultant  percentages,  i.e.,  25  per  cent  to  the  sul- 
phur, 33.5  per  cent  to  the  copper,  and  31.5  per  cent  to  the  iron.  The  grade 
of  the  matte  in  copper  is  then  33.5  per  cent, 


CALCULATION  OF  CHARGE 


399 


CHARGE-SHEET.     REVERBERATORY   SMELTING. 


Dry 
Wt 

Cu. 

SiO». 

Fe. 

CaO. 

S. 

=25% 
per  cent. 

=  33.5% 
=  31.5% 
90.0% 

Per           TK 

Cent.       Lb" 

Per      T. 
Cent.    Lb- 

Per 

Cent. 

Lb. 

Per 

Cent. 

Lb. 

Per 

Cent. 

Lb. 

Roasted  ore.  .  .  .     2000 
Limestone  200 

Cu  ii 

6       120 

20      400 
4           8 

20 

1 

»tte  = 

slag  = 
slag  = 

400 
2 

2 
50 

40 
100 

7 

=42 
=  12 

140 

120 
slag           5 

408 
For  m 

For 
FeO  for 

402 
108 

140 
Volatilized 

In  slag 

In  matte 
Wt.  of  matte 

Cu  and  Fe 
Cu 

Fe 

140 
54 

Cu  in  matte    115 

Slag 
SiOt  =408  =39.8  per  cent. 
FeO  =378  =36.6  per  cent. 
CaO  =  140  =  KJ.6  per  cent. 

926  =90.0  per  cent. 

294 
378 

= 
= 

86 
344 
65 

223 
115 

108 

CHAPTER    XXXI 
CONVERTING    COPPER    MATTE 

Principle  of  the  Process. — This  consists  in  treating  molten  matte  in  a 
converter,  a  receptacle  lined  with  refractory  material.  Compressed  air, 
blown  through  the  molten  bath,  as  in  steel  converting,  burns  off  the  sul- 
phur as  862  and  oxidizes  the  iron  to  FeO,  this  entering  the  slag.  The 
slag  is  poured  off,  leaving  blister-copper,  and  this  in  turn  is  poured  from  the 
converter  into  molds. 

THE  COPPER  CONVERTER 

Two  types  are  in  general  use,  the  horizontal  or  barrel  type,  and  the 
vertical  or  Great-Falls  type.  Both  are  electrically  operated. 


FIG.  217. — Horizontal  Type  of  Converter. 

The  Horizontal  Converter. — Fig.  217  is  a  view  of  the  horizontal  or  bar- 
rel type,  at  first  everywhere  used.  It  consists  of  a  cylindrical  steel  shell, 
having  riding  rings  at  each  end  by  which  it  is  carried,  and  revolved  on  four 
carrying  rollers.  By  means  of  four  links  riveted  to  the  converter  top,  the 
steel  shell  can  be  lifted  from  its  stand  and  transferred  where  desired  by 
40-ton  traveling  crane  for  relining,  a  newly  lined  shell  then  being  put  in 
its  place.  To  obtain  better  access  to  the  interior  the  separate  nose  can  be 

400 


THE  COPPER  CONVERTER  401 

unbolted  and  removed.  At  the  front  is  seen  the  rectangular  wind-box, 
having  fourteen  tuyeres.  As  better  shown  in  Fig.  218,  the  air  supply 
enters  through  a  sleeve  connection  at  the  axis  of  the  shell,  and  through  a 
cast-iron  passage  to  the  wind-box.  The  motor,  through  a  worm  and  worm 
gear,  revolves  the  converter  to  any  desired  position.  Immediately  at  the 
front  in  Fig.  217  is  shown  a  band-brake  by  which  the  motor  can  be  quickly 
stopped. 

From  the  "  blast  main,"  Fig.  224  as  marked,  a  12-in.  branch  leads  to  the 
axial  line  of  the  converter,  having  there  a  sleeve.  The  air  passing  from  this 
point  curves  around  the  converter  to  the  wind  box  where  are  the 
tuyeres,  and  below  them  the  "  puncher's  platform."  A  hinged  platform  at 


FIG.  218. — Great  Falls  or  Upright  Type  Converter. 

the  front  is  used  when  spent-accretions  are  to  be  cut  away  or  for  throw- 
ing in  cold  matte,  etc. 

The  Smith-Pierce  is  another  type  of  horizontal  converter,  basic  lined, 
and  in  successful  use  at  several  copper  plants.  They  are  made  up  to  13  ft. 
in  diameter  by  30  ft.  long.  (See  Fig.  220.) 

The  Upright  Converter. — This,  called  also  the  Great^Falls  type,  since 
it  was  there  developed,  is  built  in  sizes  up  to  20  ft.  diameter,  though  the 
12-ft.  converter  is  common.  As  shown  in  Figs.  218  and  219,  it  is  cylindiical 
in  plan  with  a  tapering  bottom  and  top.  The  converter-top  comes  apart 
just  above  the  trunnions,  so  that,  when  the  converter  has  been  removed 
for  relining,  this  may  be  unbolted  and  lifted  off.  Riding-rings  are  bolted 
to  the  shell,  and  are  carried  on  rollers  as  in  the  horizontal  type.  The 
wind-box  occupies  the  front  half  of  the  circumference,  and  its  connection 
to  the  blast-main,  also  the  control  valves,  is  plainly  shown.  There  are 


402 


"CONVERTING  COPPER-MATTE 


twenty-four  individual  tuyeres  branching  from  the  wind-box,  any  one  of 
which  can  be  separately  removed  if  desired.  The  hole  through  the  lining 
for  each  is  1J  to  1J  in.  In  the  view,  Fig.  218,  can  be  seen  the  end  of  the 
worm  drive,  but  not  the  electric  motor. 

THE  CONVERTER  LINING 


Both    acid    and    basic    linings    have 


FIG.  219. — Sections  of  12-ft.  Basic  Converter. 


The  Basic   Lining. — The   usual   practice    is 
with    magnesite    brick    to    a   thickness   varying 


been  used,  the  first  exclu- 
sively for  many  years.  It 
was  contended  that  the 
acid  lining  was  necessary 
in  order  to  furnish  silica 
for  the  slag  necessarily 
produced  as  the  iron  was 
oxidized.  An  acid  lining 
would  last  from  seven  to 
nine  heats,  or  much  less 
than  twenty-four  hours, 
and  then  had  to  be  re- 
moved for  relining.  This 
was  done,  using  a  ganister 
of  85  per  cent  silica  and 
15  per  cent  clay;  the  con- 
verter was  then  dried  and 
heated  for  re-use.  The 
lower  the  grade  of  the 
matte,  that  is  the  higher 
it  was  in  iron,  the  shorter 
the  life  of  the  lining,  so 
that  the  converting  of 
low-grade  mattes  was  pro- 
hibitory. 

It  was  found  that  by 
supplying  silica  to  the 
charge  directly  upon  the 
surface  of  the  molten  bath 
the  operation  could  be 
carried  forward  with  a 
basic  lining,  and  so,  since 
1911,  the  acid  lining  has 
been  given  up  in  favor  of 
the  basic. 

to    line    the    converter 
from   24   in.    at    the 


THE  COPPER  CONVERTER  403 

tuyeres  to  9  in.  elsewhere.  But  the  temperature  in  the  converter  is  con- 
tinually varying;  the  lining  is  hot  during  the  blowing  period,  and  cools 
during  pouring  or  recharging.  These  variations  in  temperature  cause  the 
lining  to  crack  and  spall  off,  also  there  is  the  mechanical  wear  of  the  charge, 
contributing  to  shortening  its  life.  This  led  to  the  idea  of  forming  a  pro- 
tective coating.  In  blowing  the  initial  charge  of  matte  without  silica  and 
at  a  moderate  temperature,  it  was  found  that  above  the  iron  oxide  neces- 
sary to  combine  with  any  silica  present  there  was  an  excess  of  iron,  which 
under  the  oxidizing  action  of  the  blast,  was  converted  to  magnetite 
(FesCU).  At  about  1200°  C.  this  magnetite  becomes  mushy  and  attaches 
itself  to  the  lining,  forming  the  needed  protective  coating,  which  can 
by  control  of  the  temperature  be  added  to  at  will.  With  proper  opera- 
tion and  care,  the  lining,  thus  protected,  should  last  indefinitely.  It 
has  the  farther  advantage  that  low-grade  mattes  can  be  readily  treated. 


FIG.  220.— Pierce-Smith  Horizontal  Converter. 

OPERATION  OF  THE  BASIC  CONVERTER 

The  operation  of  the  horizontal  converter  may  be  thus  described. 
The  initial  charge  is  60  tons  of  matte  to  which  is  added  10  per 
cent  of  dried  quartz  for  fluxing.  The  blast  is  now  turned  on  for  thirty 
to  thirty-five  minutes  and  the  charge  begins  to  heat  up  owing  to  the  active 
oxidation  of  the  FeS  of  the  matte.  The  converter  then  is  tilted  to  pour 
slag,  the  blast  being  at  the  same  time  shut  off.  With  the  completion  of 
the  pouring  a  ladleful  or  6  tons  is  introduced  and  3  tons  of  silicious  ore 
added.  Another  blow  then  begins.  This  cycle  of  blowing  and  addition 
of  matte  and  silicious  flux  is  continued  until  70  to  80  tons  of  blister-copper 
has  accumulated.  It  means  the  charging  of  300  to  400  tons  of  matte,  and 


404  CONVERTING  COPPER-MATTE 

a  period  of  thirty  to  fifty  hours  of  blowing  time  according  to  the  grade 
of  the  matte.  The  molten  copper  is  then  poured  into  large,  hot,  lined 
ladles  and  transferred  to  the  casting-furnace  or  directly  to  the  casting- 
machine. 

Operating  Precautions. — As  converting  is  primarily  an  oxidation 
process,  the  speed  of  working  depends  on  the  speed  of  blowing.  Hence 
the  tuyeres,  which  tend  to  slag  over  if  untouched,  should  be  kept  well 
punched  by  .the  insertion  of  a  punching  bar,  as  is  done  in  blast-furnace 
smelting,  except  that  in  converter  practice  this  is  done  every  few  minutes. 
In  this  way  the  tuyeres  are  kept  open,  bright,  and  in  working  order.  The 
temperature  of  the  converter  can  be  regulated  by  the  addition  of  cold 
matte,  ore  and  the  rich  sweepings  that  get  spilled,  and  that  accumulate 
during  operations.  An  addition  of  a  small  amount  of  hot  matte,  about  ten 
minutes  before  the  charge  is  finished,  will  ensure  hot  copper  when  pouring. 

Introduction  of  Silicious  Ore  to  the  Converter. — This  has  been  done 
largely  by  using  charging  boats  or  trays  that  will  hold  1  or  2  tons  of  mate- 
rial. These  are  provided  with  chains  by  which  they  are  lifted  and  handled. 

CHEMICAL  REACTIONS  OF  THE  CONVERTER 

The  Slagging  Period. — Referring  to  operations  in  a  12-ft.  converter, 
the  matte  may  be  taken  as  containing  Cu,  43  per  cent;  Fe,  29  per  cent,  and 
S,  14  per  cent;  corresponding  to  Cu2S  and  FeS  with  a  little  Fe304.  The 
period  begins  when  the  first  ladleful  of  matte  has  been  poured  in  and  the 
blast  turned  on.  The  heat  at  first  drives  off  elemental  sulphur  according 
to  the  reaction: 

(1)  5FeS+heat  =  Fe5S4+S. 

(2)  Fe5S4+  14O  =  2FeO+Fe3O4+4S02. 

Here  is  revealed  the  source  of  the  material  to  furnish  a  magnetite  lining. 
But  a  subsidiary  reaction  then  takes  place  by  which  this  magnetite  is 
reduced  to  FeO,  then  entering  the  slag: 

(3)  Fe3O4+Fe5 =4FeO+4FeS. 

The  principal  reaction  now  becomes:  * 

(4)  Fe5S4+ 140  =  2FeO+Fe304+4S02. 

In  this  vigorous  reaction  much  sulphur  dioxide  is  evolved,  the  magnetite 
becomes  reduced,  according  to  reaction  (3)  into  FeO,  and  this  together  with 
that  resulting  from  reaction  (2),  also  enters  the  slag.  It  is  thus  s"een  how 
necessary  it  is  at  this  stage  to  add  silicious  ore  to  unite  with  the  FeO. 
The  reactions,  at  first  slow,  rapidly  increase  until  in  about  forty-five  min- 
utes they  are  completed;  much  of  the  sulphur  is  gone  and  most  of  the  iron 


CHEMICAL  REACTIONS  OF  THE  CONVERTER 


405 


slagged,  and  the  converter  contents  brought  to  the  stage  of  white  metal. 
The  end  of  the  stage  is  known  by  the  appearance  of  the  issuing  flame,  the 
greenish  border,  at  first  seen,  changing  to  a  pale  permanent  blue.  Pieces 
of  matte  are  thrown  into  the  converter  if  it  is  wished  to  make  the  charge 
hotter,  while  sweepings  from  around  the  converter,  rich  in  copper,  when 
added  tend  to  make  it  cooler.  At  this  time  the  converter  is  turned  down 
and  the  slag  poured  into  the  slag-ladle.  To  tell  when  the  matte  begins  to 
escape,  "  the  skimmer  "  passes  a  rabble  through  the  flowing  stream  and 
can  thus  determine  the  presence  of  drops  of  matte,  whereupon  the  converter 
is  returned  to  blowing  position.  This  converter  slag,  containing  1.5  to 
2  per  cent  copper  and  about  0.5  to  0.1  oz.  silver  per  ton,  is  sent  to  the  blast- 
furnace for  recovery  of  its  metal  contents. 

Conversion  of  White  Metal  to  Copper. — At  the  beginning  of  this  blow 
there  is  but  little  iron  left,  and  the  matte  has  been  brought  to  the  stage  of 
white  metal  of  75  per  cent  Cu,  and  we  have: 

(5) 
(6) 


=  2Cu+SO2, 
4Cu2S+9O  =  6Cu+Cu2O+4SO2, 


Van  Liew 


FIG.  131.  —  Elimination  of  Impurities  in  Converting. 

with  a  further  abundant  evolution  of  SO2  and  the  formation  of  blister- 
copper.  The  Cu2O  of  reaction  (6)  is  soon  reduced  to  Cu  as  per  reaction  (5). 
As  far  as  the  agitation  of  the  blast  permits,  the  molten  contents  separate 
into  layers,  an  increasing  layer  of  white  metal  above,  and  slag  on  top. 
The  blast  enters  the  bath  horizontally  6  to  12  in.  above  the  bottom,  and 
blows  largely  through  the  white-metal  layer.  The  escaping  flame  is  white, 
gradually  changing  to  rose-red  and  finally  to  a  brownish  red.  It  decreases, 
until  at  last  there  is  but  a  brick-red  flickering.  The  identification  of  the 
finish  needs  care  and  experience.  If  carried  too  far  we  have  over-blown 
copper.  The  converter  is  now  turned  down,  the  blast  being  at  the  same 
time  shut-off,  and  the  blister  is  poured,  either  directly  into  molds,  or  into 
a  ladle,  to  be  taken  to  a  tilting-furnace.  The  converter  is  then  turned 
back  to  receiving  position  for  treatment  of  a  new  charge.  The  air  is 
supplied  at  a  pressure  of  12  to  15  Ib.  per  square  inch,  using  150,000  cu.  ft.  of 


406 


CONVERTING  COPPER-MATTE 


air  per  ton  of  blister  produced.  It  takes  four  hours  for  a  cycle  of  operations, 
and  in  this  time  25  tons  of  matte  are  treated,  yielding  10  tons  of  blister- 
copper. 

Loss  in  Converting. — The  escaping  gases  contain  nitrogen,  sulphur 
dioxide,  and  traces  of  volatilized  metals.  The  loss  of  gold  is  small;  that 
of  silver  depends  upon  the  amount  of  volatile  metals.  In  the  flue-dust  in 
one  case  there  was  an  average  of  40  oz.  per  ton.  The  loss  in  converting 
may  be  given  at  1  to  1.5  per  cent  of  the  copper  and  2  to  2.5  per  cent  of 
the  silver.  When  treating  leady  matte  from  a  silver-lead  furnace,  this 

loss  of  silver  may  be  serious, 
amounting  to  33  to  40  per  cent. 
At  Tooele,  Utah,  where  the  lead 
and  zinc  fumes  are  caught  in  a 
bag-house,  the  silver  is  saved. 

We  give,  in  Fig.  221,  a  graphic 
chart  showing  the  losses  in  an 
acid-lined  converter  where  the 
period  of  the  blow  was  seventy 
minutes.  It  shows  that  after  ten 
minutes  the  1.2  per  cent  of  zinc 
is  gradually  burned  off.  Anti- 
mony and  arsenic,  both  present 
to  the  extent  of  0.25  per  cent  in 
one  case  and  to  0.15  per  cent 
in  another,  are  well  eliminated 
toward  the  end. 

Crane  Ladle.— Fig.  222  is  a 
view  of  a  steel  ladle  used  for 
transferring  matte  or  blister-cop- 
per by  means  of  the  traveling 
crane  from  the  forehearth  of  a 
blast-furnace  or  from  a  reverberatory  furnace.  When  in  position  it  is 
tipped  by  an  auxiliary  hoist  of  the  crane  which  hooks  into  the  eye  shown 
at  the  left  side  of  the  ladle.  The  ladle  is  plastered  on  the  inside  with  a 
coating  of  clayey  loam,  which  is  dried  out  and  heated  before  using.  - 


FIG.  222.— Crane  Ladle. 


BLAST-FURNACE  SMELTING  AND  CONVERTING  PLANT 

Fig.  224  represents  the  plan  of  a  blast-furnace  plant  with  a  two-stand 
converter-plant  attached,  the  size  of  the  latter  being  indicated  by  the 
number  of  stands  or  stalls  in  which  copper  matte  can  be  blown.  In  the 
elevation,  Fig.  223,  the  receiving  track  for  coke  is  shown  at  the  extreme 
right  of  the  illustration  unloading  into  the  coke  bins  beneath.  The 


BLAST-FURNACE  AND  CONVERTER  BUILDING 


407 


408 


CONVERTING  COPPER-MATTE 


ore-bins  with  the  inclined  bottoms  are  shown  to  be  on  the  same  level. 
In  the  furnace  building  there  are  two  matting  blast-furnaces,  each  42  by 
144  in.  at  the  tuyere  level  and  each  having  a  settler  or  fore-hearth  10  ft. 


diameter.     Blast  is  furnished   to  the  furnaces  from  a  power-house,  not 
shown. 

In  Fig.  223  is  shown  the  semi-elliptical  flue  3  ft.  by  7  ft.  high,  leading 


THE  COTTRELL  ELECTROSTATIC  TREATER 


409 


from  the  blast-furnaces.  This  crosses  the  near  end  of  the  furnace  building, 
as  indicated  by  the  dotted  lines  in  Fig.  224.  It  is  connected  to  a  dust- 
chamber  which  leads  to  a  stack.  The  slag  is  taken  away  over  an  electric- 
trolley  system  entering  the  building.  The  slag-cars  are  brought  close  to 
the  settlers  to  receive  the  flowing  slag  while  the  matte,  as  needed,  is  tapped 
from  a  lower  tap-hole  into  the  steel  ladle  for  transferring  to  the  converters. 
A  platform  elevator,  at  the  end  of  the  furnace  building,  elevates  slag  and 


FIG.  225.— Cottrell  Treater. 


other  material  from  the  floor  of  the  converter  building  to  the  charge-floor 
for  the  blast-furnace. 

The  converter  building  has  at  one  end  the  lining  floor,  and  is  com- 
manded from  end  to  end,  by  a  40-ton  electric  traveling  crane  which  serves 
to  handle  the  converters,  to  supply  them  with  matte,  to  take  away  the 
slag,  and  to  handle  all  materials  for  and  from  the  converters.  For  each 
stand  there  should  be  an  extra  shell,  or  four  in  all.  The  escaping  gases 
from  the  converters  are  received  in  a  hood  attached  to  a  dust  chamber 


410 


CONVERTING  COPPER-MATTE 


so  that   the  particles  of   matte   blown   out   by  the  blast   are   collected. 
From  the  hood  a  flue  connects  with  a  dust  chamber  and  this  with  a  stack. 
At  one  end  of  the  furnace  building  (see  Fig.  224)  is  the  mill  where  the 
lining  material  is  prepared  for  the  converter  (now  no  longer  used) . 

ELECTROSTATIC  RECOVERY  OF  COPPER  BLAST-FURNACE  AND 
CONVERTER  DUST 

This,  called  also  the  Cottrell  process,  has  come  into  use  since  it  is 
possible  by  it  to  remove  the  dust  and  fume  from  the  hot  gases  produced 
in  roasting  fine  ore,  or  the  fumes  escaping  at  the  converter-plant  where,  due 
to  the  high  temperature,  a  bag-house  could  not  be  used. 

In  principle,  the  fume  and  dust  in 
suspension  flow  upward  through  a 
vertical  tube  5  to  10  in.  in  diameter 
by,  say,  15  ft.  long,  as  shown  in  Fig. 
226.  No.  10  insulated  copper  wire, 
suspended  axially,  takes  a  high-voltage 
undirectional  current  of  25,000  to 
60,000  volts  or  more.  The  current, 
passing  from  the  wire  to  the  inner  sur- 
face of  the  tube,  electrifies  the  fume 
or  dust  particles  negatively,  and  these 
are  repelled  to  the  inner  surface  of 
the  tube,  forming  a  coating  upon  it. 
This  coating  is  occasionally  removed 
by  jarring  it  off  into  a  hopper  below, 
whence  it  is  removed  by  a  spiral- 
screw  conveyor. 

Another  method,  called  the  plate  system,  consists  in  having  vertical 
corrugated  plates  about  10  in.  apart  and  at  10  in.  intervals  chains  of  J 
in.  diameter  suspended.  The  current  down  those  chains  acts  in  a  similar 
way,  repelling  the  dust  particles  to  the  surface  of  the  plates.  A  later  varia- 
tion of  this  consists  in  having  rods  passed  horizontally  equidistant  between 
the  plates  to  carry  the  current.  A  chamber  full  of  pipes  or  plates  with  the 
rods  thus  insulated  is  called  a  Cottrell  treater  unit.  The  insulate^  wires 
are  connected  in  one  and  receive  the  high-pressure  current. 

THE  WORKS  OF  THE  INTERNATIONAL  SMELTING  CO. 

This  comprises  two  separate  installations,  a  roasting  and  drying  plant 
and  a  reverberatory  smelting  and  converting  plant.  The  dried  or  roasted 
product  of  the  first  being  smelted  at  the  second,  we  have,  therefore: 

(1)  A  roaster  or  dryer  plant  with  a.  Cottrell  treater. 


FIG.  226. — Diagram  of  Treater. 


ROASTING  PLANT 


411 


(2)  A  smelting  and  converter-plant  where  the  converter  dust  is  caught 
in  a  Cottrell  treater. 

At  the  dryers  or  roasters  (see  Fig.  227)  are  trippers  that  regularly  feed 
to  five-hearth  Wedge  roasters.  Beneath  are  two  rows  of  calcine  hoppers, 
so  placed  that  the  calcine  can  be  drawn  off  into  cars  that  take  it  away  to  the 
smelting  works.  In  Fig.  236 
is  shown  one  of  the  two  fire- 
boxes which  heat  the  roasters, 
two  and  two,  respectively. 
The  roaster  arms  of  the  fur- 
nace are  cooled  by  air  and  the 
delivery  of  this  air  is  by  an 
underground  pipe  marked 
"cooling  air"  in  Fig.  236. 
Above  and  between  the 
roasters  is  the  gas  flue, 
which  receives  the  branch 
pipes  from  the  roaster.  From 
the  top  of  the  gas  flue  are 
pipes  that  branch  right  and 
left  to  the  respective  gas 
treaters  for  each  roaster. 
Each  treater  has  thirty-six 
pipes  12  in.  diameter  by  15 
ft.  long  with  an  upward  flow 
of  gases  through  them.  Here 
the  dust  is  separated,  and  the 
gas  escapes  by  stacks  or 
chimneys,  three  to  each 


ELEVATION  A-A 


FIG.  227.— End  Elevation  of  Roaster  Plant. 


treater,  high  above  the  build- 
ing. 

The  Roaster  Plant.— This 
roasting,    or    rather    drying 

plant  of  the  company  is  for  the  drying  of  a  flotation  concentrate,  to  which 
is  added  a  small  amount  of  sulphide  ore  hi  order  to  produce  the  needed 
proportion  of  matte  hi  the  subsequent  smelting.  The  material  is  so  fine 
that  a  Cottrell  installation  had  to  be  added  in  order  to  prevent  excessive 
loss  of  flue  dust. 

Figs.  227  and  228  are  two  views  of  the  roaster  building,  which  con- 
tains five  Wedge  roasters,  each  roaster  having  its  own  treater.  In  the 
other  end  of  the  building  is  the  "  electrical  machinery  room  "  con  taming 
the  switch  boards,  generator-exciter  sets  and  the  motor-generator- 
transformer  sets. 


412 


CONVERTING  COPPER-MATTE 


Here  are  the  elevating  and  conveying  belts  for  delivery  of  the  flotation 
concentrates  to  the  furnaces.  There  is  so  little  sulphur  in  this  product  that 
there  is  no  need  to  roast  it,  and  the  wet  material  is  simply  dried  before 
being  sent  to  the  reverberatory  for  smelting.  Were  it  needed,  roasting 
could  be  added.  The  material  being  so  fine  much  flue  dust  is  necessarily 
made. 

The  Smelting  and  Converting  Plant. — Fig.  229  is  a  plan  of  this  smelting 
plant  for  the  treatment  of  a  flotation  concentrate,  to  which  has  been  added 


ELEVATION  C-C 


ELEVATION  D-D 


WWM0P 

ELEVATION  E-B 


FIG.  228.— Side  Elevation  of  Roaster  Plant. 


\ 


a  small  amount  of  sulphide  in  order  to  produce  the  needed  proportion  of 
matte.  The  plant  consists  of  the  converter  house,  its  Cottrell  treater  and 
stack  at  the  right,  the  reverberatory  building  at  the  center,  and  the  boiler 
house  for  the  waste-heat  boiler  with  their  flue  and  stack  at  the  left,  also 
auxiliary  equipment  of  the  plant  above. 

There  are  three  large  reverberatory  furnaces  (oil-fired),  over  which 
transversely  run  a  double  line  of  tracks  which  bring  in  the  calcines  from  the 


REVERBERATORY  SMELTING  AND  CONVERTING  PLANT        413 

dryer  and  roaster  building,  some  distance  away.  The  firing  end  of  the 
reverberatories  adjoins  the  converter  house  set  at  a  lower  level,  so  that  the 
matte  ladles  can  be  set  low  enough  to  take  the  matte  when  tapped.  At 
the  front  end  are  seen  the  slag  cars  on  a  sunken  track.  As  fast  as 
filled  they  are  removed  by  a  locomotive. 

Waste-heat  Boilers. — Over  this  sunken  track  is  the  heater  flue,  which 


Converters  Slag 
Launder 
5,  12  ft.  Great  Falls 
Type  Converters 


FIG.  229. — Plan  of  Reverberatory  Smelting  and  Converting  Plant. 

takes  the  gases  from  the  three  furnaces.  From  this  flue  there  are  six 
branch  flues  to  the  six  waste-heat  boilers,  and  each  of  these  in  turn 
branches  to  the  reverberatory  flue,  this  latter  leading  to  the  main  stack 
300  ft.  high  by  25  ft.  diameter. 

The  Cottrell  Treater  for  the  Converters.— In  the  converter  house  are 
five  converter  stands,  each  arranged  as  in  Fig.  229,  the  goose-neck  branch 
pipe  leading  to  a  dust  bin,  the  hopper  of  a  Cottrell  treater  where  a  part 


414  CONVERTING  COPPER-MATTE 

of  the  dust  settles  out.  The  rest  of  it  is  taken  out  by  a  converter  Cottrell 
system  so  that  the  valuable  dust  is  quite  recovered.  The  gases  pass  away 
by  the  converter  stack.  Beneath  the  dust  bin  is  a  dust  track  and  paral- 
lel to  it  a  copper  bullion  track  for  the  removal  of  flue  dust  and  the 
blister  copper  respectively. 

The  Tilting  Furnaces. — There  are  two  of  these  Nos.  1  and  2  of  the  plan, 
Fig.  229,  so  that  one  furnace  is  filling  while  the  other  is  pouring.  They 
resemble  a  large  horizontal  converter  and  receive  a  number  of  ladlesful  of 
blister  as  this  is  made  at  the  converters.  When  full,  the  tilting  furnace  is 
poled  to  make  a  smooth  ingot  and  is  poured  into  the  casting  machine 
adjoining,  see  Figs.  242  and  243.  The  ingots,  as  they  fall  into  the  water 
bosh  of  the  casting  machine,  are  there  cooled,  then,  by  an  endless  chain 
are  raised  and  delivered  into  the  casting' shed  for  weighing  and  shipping 
away  by  the  copper  bullion  track. 

Other  Equipment. — The  converter  air  main  brings  in  the  air  from 
the  compressor  in  the  power  house  (not  shown)  at  a  pressure  of  15 
Ib.  per  square  inch  for  use  at  the  converters.  A  battery  of  eight 
oil  tanks  supplies  the  reverberatory  furnaces  through  an  8-in.  main. 
Near  these  tanks  is  seen  the  installation  of  "  mud  bins  "  where  clay 
is  stored  for  mixing  with  crushed  silica  ore  in  the  silica  bin.  These 
are  mixed  to  form  ganister  in  a  Carlin  mill  for  use  in  lining  the  con- 
verters. They  are  now  little  used  under  conditions  of  modern  practice. 
The  skull  breaker  consists  of  a  strong  grated  hopper  into  which  the  skulls 
or  shells,  that  form  like  a  lining  on  the  interior  of  the  ladles,  are  broken.  A 
weight  lifted  by  the  traveling  crane  is  let  fall  upon  these  skulls,  breaking 
them  to  a  size  for  convenient  handling  so  that  the  pieces  can  be  charged 
into  the  furnace  for  melting  down. 

COSTS  OF  A  PROPOSED  PLANT  AND  OPERATION 

For  the  year  1919  we  give  these  costs  for  a  proposed  plant  for  the 
Consolidated  Copper  Mines  Co.  to  be  built  at  Kimberly,  Nev.  For  this 
Frederick  Laist,  in  charge  of  the  works  at  Anaconda,  Mont.,  who 
planned  it,  prescribes  a  single  reverberatory-furnace  plant  of  445  tons 
daily  capacity,  using  a  charge  of  these  items: 

(1)  133  tons  of  concentrates  of  the  composition  Cu,  18  per  cent;    SiC>2, 
20  per  cent;  Fe,  25  per  cent,  yielded  from  2000  tons  daily  of  "  porphyry 
ore  "  of  1.2  Cu. 

(2)  37  tons  concentrates  of  the  composition  Cu,  10.5  percent;  SiO2, 
15  per  cent;  Fe,  32  per  cent,  and  S,  34  per  cent,  made  from  150   tons 
"  sulphide  "  ore  containing  2.9  copper. 

(3)  150  tons  of  oxidized  ore  containing  7.5  per  cent  copper. 

(4)  125  tons  of  fluxing  materials. 


ESTIMATED  COST  OF  REDUCTION  WORKS  415 

It  is  assumed  that  these  ores  can  be  delivered  to  the  reduction  works 
at  a  cost  of  $1.10  for  the  porphyry  ore,  $5  for  the  sulphide  ore,  and  $10  for 
the  oxidized  ore. 

The  reduction  works  needed  for  treating  this  quantity  of  ore  consists 
of  a  concentrating  plant  having  a  capacity  for  2000  tons  of  porphyry  ore 
and  150  tons  of  sulphide  ore  also  a  power  plant  capable  of  generating  3000 
kw.  which  will  furnish  the  power  needed  for  water  supply,  operating  the 
mining,  concentrating  and  smelting  plants. 

The  cost  of  the  water  supply  is  high,  since  it  must  be  pumped  to  a  total 
height  of  967  ft.,  or  against  a  total  head  of  1200  ft.  and  a  distance  of  13 
miles. 

The  concentrator  is  to  be  constructed  in  two  sections  and  provided 
with  spare  grinding  mills  and  flotation  machines,  so  that  a  breakdown  of 
one  of  these  will  not  affect  operations. 

The  smelting  plant  mill  comprises  four  20  ft.,  seven-hearth  Wedge 
roasters;  one  100  ft.  by  20  ft.  reverberatory  furnace  equipped  with  waste 
heat  boilers  and  a  converting  department  containing  two  12-ft.  upright 
converters.  The  smelter  building  should  be  made  of  size  to  accommodate 
another  reverberatory  furnace.  Such  extension  would  cost  $50,000, 
while  a  second  furnace  would  cost  $125,000  more. 

ESTIMATED  COST  OF  REDUCTION  WORKS  AS  OUTLINED  HEREWITH 

Assume  daily  treatment — Porphyry  ore 2,000  tons 

"  Sulphide  "  ore 150  tons 

"  Oxidized  "  ore 150  tons 

Water  supply  1000  to  1500  gal.  per  minute $305,740 

Crushing  plant • 100,000 

Concentrating  plant 450,000 

Power  plants— 3000  kw 450,000 

Drying  plant 200,000 

Reverberatory  plant 350,000 

Converting  plant 150,000 

Bins,  rolling  stock,  shops,  houses  and  miscellaneous 300,000 


$2,305,740 
Engineering,  drafting  and  contingency  (say) 194,260 


Total $2,500,000 

This  estimate  is  based  on  present  cost  of  supplies  and  labor  and  is  con- 
sidered conservative. 

The  plant  can  be  enlarged  at  any  time,  without  interfering  with  opera- 
tions, to  5000  tons  of  porphyry  and  250  tons  of  oxidized  ore,  at  an  addi- 
tional expense  of  about  $1,400,000. 


416  CONVERTING  COPPER-MATTE 


OPERATING  EXPENSES 

Assume  smelter  recovery  at  95  per  cent,  operating  costs  are  estimated  as  follows. 

Power — $0.01  per  kw.-hr. 

Concentrating — $0.85  per  ton. 

Drying — $0.45  per  ton. 

Reverberatory  smelting — $2.35  per  ton. 

Converting  and  casting — $10  per  ton  of  copper. 

TOTAL  SMELTING  EXPENSES 

Crushing  275  tons  ore,  flux  and  secondaries  at  15  cents $41 . 20 

Roasting  445  tons  at  $0.45 200 . 00 

Reverberatory  smelting  445  tons  at  $2.35 1,045 . 00 

Converting  37  tons  Cu  at  $7.50 . . ; . 278.00 

Casting  and  loading  27  tons  at  $2.50 92 . 60 


$1,656.80 
Add  10  per  cent  for  miscellaneous *.  .  . , 165 . 68 


$1,822.48 

Cost  of  smelting  per  ton  ore  concentrate  mixture 5 . 70 

Cost  of  smelting  per  Ib.  of  copper  produced 0 . 0246 

Cost  of  concentrating  per  Ib.  of  copper  produced 0.0247 

Treatment  of  Porphyry  Ores  Alone. — In  order  to  convey  an  idea 
of  the  value  of  the  porphyry  ores  alone,  without  admixture  of  high- 
grade  "  oxidized  "  ore,  the  following  estimate  is  submitted.  The  complete 
treatment  of  porphyry  ore  alone  locally  would  scarcely  be  feasible  on  a 
scale  of  much  less  than  5000  tons  per  day.  This  quantity  of  ore  would, 
however,  yield  approximately  300  tons  of  concentrates,  which  would 
make  the  total  amount  of  material  to  be  smelted,  including  fluxes  and 
secondaries,  about  425  tons  per  day,  which  would  be  an  economical  opera- 
tion for  one  reverberatory  furnace. 

The  cost  of  the  operation  would  be  about  as  follows : 
Assume  recovery  in  bullion  of  80  per  cent  of  1.4  per  cent  Cu  =  22.4  Ib. 
per  ton. 

COST   OF  TREATMENT  PORPHYRY   ORE  ALONE— 5000  TONS   PER   DAY 

Per  Lb.  Cu 

Mining  at  $1.10  per  ton  of  ore |0 . 0491 

Concentrating  at  $0.85  per  ton  of  ore .  0379 

Smelting  at  $6  per  ton  of  concentrates .0179 

Freight  at  $16.60  per  ton  copper .  0083 

Refining  at  $22  per  ton  copper .  01 10 


Gross  cost. .  .  $0. 1242 

Credit  for  gold  and  silver .  0050 


Net  cost $0 . 1192 


CHAPTER  XXXII 
THE  HYDROMETALLURGY  OF  COPPER 

PRINCIPLES  OF  THE  HYDROMETALLURGY  OF  COPPER 

The  wet  methods  of  extracting  copper  from  cupiferous  ore  consist  in 
obtaining  the  copper  from  the  crushed  and  perhaps  roasted  ore,  in  water 
solution,  either  with  or  without  the  aid  of  other  solvents  such  as  a  solution 
of  ferric  oxide  or  of  sulphuric  or  hydrochloric  acids.  The  copper  must  be 
in  combination  with  elements  that  will  permit  it  to  dissolve  in  the  solvents 
used.  Thus,  metallic  copper  would  not  dissolve  in  sulphuric  acid,  and 
chrysacolla  is  difficultly  soluble.  From  the  clear  decanted  or  filtered 
copper-bearing  solution  the  metal  may  be  precipitated  electrolytically,  or 
with  scrap-iron  or  with  lime.  The  resultant  "  precipitate  "  is  then  melted 
and  refined. 

Available  Copper  Ores. — Copper  has  been  extracted  profitably  from 
suitable  ore  of  as  low  grade  as  0.5  to  1.5  per  cent  copper  when  the  condi- 
tions of  an  abundant  and  easily  exploited  supply  and  cheap  labor  pre- 
vailed. Such  bodies  of  ore,  much  of  it  in  oxidized  form,  occur  throughout 
the  world,  often  of  too  low  grade  to  be  treated  by  smelting,  or  too  difficult 
of  access.  Besides  this  there  are  huge  dumps,  being  the  tailings  of  con- 
centrating mills.  Ore  containing  the  copper  as  oxide,  carbonate  or  sul- 
phate is  best  suited  to  extraction,  but  if  containing  lime,  magnesia,  ferrous 
oxide,  or  manganese  oxide,  it  is  less  desirable.  Copper-bearing  sulphides 
may  be  profitably  treated  for  the  extraction  of  the  metal,  the  ore  being 
oxidized  by  weathering  until  the  sulphide  has  been  changed  into  a  sulphate 
soluble  in  water.  Sulphide  ores  may  be  roasted  with  salt  to  bring  them 
into  form  of  a  chloride,  which  then  is  extracted  with  brine  solution. 

Advantages  of  Leaching. — It  would  seem  that  for  low-grade  ores 
leaching  should  be  superior  to  other  methods,  since  the  worthless  gangue, 
which  is  the  largest  constituent  of  the  ore,  remains  untouched,  and  the* 
solvent  acts  on  the  relatively  small  quantity  of  valuable  metal.  Particu- 
larly does  this  seem  to  be  true  of  silicious  ores,  the  silica  of  which  in  no 
way  interferes  with  the  leaching,  while  they  are  expensive  to  smelt. 


417 


418  THE  HYDROMETALLURGY  OF  COPPER 

EXTRACTION  OF  COPPER  BY  NATURAL  OR  WEATHERING  METHODS 

(1)  By  direct  treatment  of  the  raw  or  crude  ore  (Rio  Tinto  process). 

(2)  By  treatment  of  the  ore  which  has  been  subjected  to  a  preliminary 
roast  in  heaps  (Shannon  Copper  Co.  process). 

(1)     THE  RIO  TINTO  PROCESS 

Outline  of  Process. — The  copper  in  the  sulphide  ore  is  brought  into 
soluble  form  as  sulphate,  and  from  the  filtrate  the  copper  is  precipitated 
by  means  of  scrap  iron.  The  copper-bearing  pyrite  is  made  into  large 
flat-topped  heaps  which  are  oxidized  by  means  of  a  regulated  supply  of 
water  and  air,  and  when  the  copper  has  been  changed  into  sulphate,  the 
material  is  leached  with  water  to  extract  this.  The  clear  solution  is  con- 
ducted to  tanks  filled  with  pig  iron  where  the  copper  is  precipitated. 

When  the  copper  in  the  ore  occurs  as  chalcopyrite  (CuFeS2)  or  as  covel- 
lite  (CuS)  oxidation  proceeds  slowly  and  imperfectly  and,  for  successful 
working,  it  should  be  in  the  form  of  chalcocite  or  copper  glance  (Cu2&) .  It 
is  because  of  the  extent  of  the  ore  bodies  and  the  cheapness  of  labor  that  the 
Rio  Tinto  process  has  been  successful.  This  ore  contains  on  an  average 
2  per  cent  copper. 

Preparation  of  Site  and  Heaps. — A  site  is  chosen  upon  impervious 
sloping  ground  for  suitably  draining  off  the  solution  as  formed.  A  clayey 
or  rocky  bottom  is  required,  or  one  properly  puddled  or  coated  with  clay  to 
render  it  impermeable.  The  heaps  may  contain  100,000  tons  of  ore  and 
are  constructed  as  follows:  On  the  ground  is  first  arranged  a  network  of 
flues  12  in.  square,  made  of  lump  ore.  Vertical  flues  or  chimneys  that  con- 
nect with  the  ground  flues  are  built  50  ft.  apart  as  the  heap  is  made.  The 
ore  is  broken  to  3  or  4  in.  diameter  and  some  of  the  lumps  are  screened  out 
for  making  the  flues,  leaving  some  fine.  The  run  of  ore  is  dumped  and 
spread  on  the  site  in  layers  until  this  is  30  ft.  high.  The  flat-top  surface, 
having  a  grade  of  one  in  300,  is  formed  into  20  ft.  squares  by  ridges  of  fine 
ore  so  as  to  ensure  distribution  of  water  within  specified  limits,  and  wash- 
ing of  the  pile  from  the  top  to  the  ground  flues  or  drains.  Launders  are 
provided  to  carry  water  to  the  heap. 

First  Operation. — As  the  heap  is  forming,  water  is  applied  to  extract 
any  already-formed  copper  sulphate.  Oxidation  starts  as  the  result  of 
the  wetting.  The  completed  and  wetted  heap  begins  to  oxidize  rapidly, 
as  shown  by  the  heat  evolved,  the  temperature  of  the  air  in  the  chimneys 
rising  to  70°  C.  As  the  heat  increases  the  ground  flues  are  closed  to  con- 
trol oxidation,  and  to  spread  the  reactions  through  the  heap.  The  sur- 
face assumes  a  brown  color,  due  to  the  dehydration  of  the  basic  ferric  salt 
that  forms,  and  heating  is  made  apparent  by  this  drying  action.  Great 
care  is  taken  to  prevent  the  heap  from  catching  fire. 


RIO  TINTO  PROCESS  419 

Chemistry  of  the  Process.  —  By  the  combined  action  of  air  and  moisture 
the  following  reactions  occur: 

(1)  FeS2+70+H20  =  FeS04+H2S04; 

that  is,  pyrite  is  oxidized  to  ferrous  sulphate  and  sulphuric  acid.  This 
ferrous  sulphate  is  readily  oxidized  to  ferric  sulphate  thus: 

(2)  2FeSO4+H2SO4+O  =  Fe2(SO4)3+H2O. 

The  thus-formed  ferric  sulphate  acts  on  chalcocite  and  changes  it  in  part 
to  copper  sulphate,  itself  reverting  to  ferrous  sulphate  according  to  this 
reaction  : 

(3)  Fe2(SO4)3  +  Cu2S  =  CuSO4  +  2FeSO4  +  CuS. 

The  cupric  sulphide,  hitherto  unaffected  is  farther  changed  as  follows: 
(4) 


Reaction  (9)  is  relatively  rapid,  and  accordingly  about  half  the  copper 
goes  into  solution  in  a  few  months.  Reaction  (10)  is  slow,  but  in  two  years, 
under  favorable  conditions,  yields  80  per  cent  of  the  remaining  half  of  the 
copper. 

Extraction.  —  When  oxidation  has  advanced  as  far  as  is  safe,  water  is 
applied  at  the  rate  of  220  gal.  per  minute  until  the  soluble  copper  salts  are 
extracted.  The  flow  is  then  stopped  and  oxidation  is  resumed,  and  is 
followed  by  renewed  washings.  After  a  year  the  top-surface  needs  retilling  ; 
the  ridges  are  arranged  where  the  squares  formerly  were  and  the  launders 
are  shifted  to  conform.  At  the  sides  of  the  heap  for  the  distance  of  some 
yards  the  ore  has  become  cemented,  and  holds  copper  salts.  These  sides 
are  dug  down  hi  terraces  to  expose  the  copper  salts  and  to  extract  them  by 
washing.  When  there  remains  but  0.3  per  cent  copper,  extraction  is  con- 
sidered complete.  This  pyrite  heap,  after  the  copper  has  been  washed 
out,  is  still  valuable  as  a  sulphur-bearing  pyrite,  and  many  tons  of  such 
washed  ore  has  been  shipped  away  to  the  sulphuric  acid  makers. 

Reduction  of  Ferric  Sulphate.  —  The  solution  that  flows  from  the  heap 
contains  ferric  sulphate,  and  to  prevent  it  from  consuming  iron  hi  the  pre- 
cipitation tanks  it  must  be  reduced  by  running  the  liquor  through  a  filter- 
bed  of  fresh  iron-pyrite  smalls  or  fines.  The  reaction  is  as  follows  : 

(5)  7Fe2(S04)3+FeS2+H2S04  =  15FeSO4+8H2SO4. 

This  filter-bed  is  retained  within  a  reservoir  formed  by  a  masonry  dam 
across  a  small  ravine.  The  liquor,  or  solution  after  percolating  the  bed, 
flows  to  a  common  settling  tank.  It  is  then  drawn  off  to  a  series  of  tanks, 
canals,  or  flumes  containing  pig  iron.  The  typical  solution  entering  the 
series  would  contain  CuO,  4  per  cent;  Fe2O3,  0.1  per  cent;  FeO,  2.0  per 


420  THE  HYDROMETALLURGY  OF  COPPER 

cent;  H2SO4,  1.0  per  cent  and  As,  0.03  per  cent.  The  presence  of  so 
much  FeO  and  H2SO4  is  due  to  the  fact  that  a  part  of  the  waste  or  bar- 
ren solution,  leaving  the  series,  is  pumped  back  and  used  for  watering  the 
heaps,  so  that  the  solution  tends  to  increase  in  these  elements. 

Precipitation.  —  The  copper-bearing  liquor  or  solution,  drawn  from  the 
filter-bed,  is  run  through  precipitation  launders  containing  pig-iron  ingots 
piled  in  open  order,  and  the  copper  is  precipitated  (replacing  the  iron  which 
dissolves)  in  the  form  of  "  cement  copper  "  or  copper  precipitate.  Fol- 
lowing the  tanks  the  solution  enters  the  canals,  flumes  or  launders,  arranged 
on  the  slope  of  a  hill  in  such  fashion  that  the  solution  may  pass  back  and 
forth  through  them  until  it  is  discharged  "  barren  "  or  free  from  copper 
from  the  end  of  the  lowest  series.  These  flumes  are  320  ft.  long,  5J  ft.  wide 
by  2J  ft.  deep.  They  begin  at  a  grade  of  2  in  1000,  and  the  final  flumes 
slope  11  in  1000.  The  rate  of  flow  is  thus  increased  and  less  pig  iron  is 
wasted.  Some  of  the  flumes  are  cut  out  from  the  flow  or  by-passed  daily, 
the  solution  meanwhile  going  through  the  remaining  ones.  Those  thus 
cut  out  are  drained,  and  all  the  pig  iron  is  removed  and  piled  beside  them, 
the  copper  attached  to  them  being  meanwhile  knocked  off  and  thrown 
back.  The  muddy  precipitate  at  the  bottom  of  the  by-passed  tanks  and 
flumes  is  removed  to  the  cleaning  and  concentrating  plant,  while  the  pig 
iron  is  piled  back,  and  the  flow  of  solution  again  directed  through  the 
flumes.  Under  the  best  conditions  there  is  needed  1.4  tons  of  pig  iron 
per  ton  of  copper  precipitated. 

The  first  reaction  in  the  tanks  is  that  between  unreduced  ferric  sulphate 
and  the  pig  iron,  thus  expressed: 

(6)  Fe2(SO4)3  +Fe  =  FeSO4. 

It  is  a  reaction  that  wastes  iron.  Precipitation  of  the  copper  is  brought 
about  by  an  electro-chemical  reaction,  viz., 

(7)  Fe+CuSO4  =  FeSO4+Cu. 

Finally  a  reaction  takes  place  between  the  free  sulphuric  acid  and  the  iron. 

(8)  Fe+H2SO4  =  FeSO4+H2. 


This  reaction  is  evinced  by  bubbles  of  hydrogen  rising  through  the  tank 
liquor.     It  means  a  further  waste  of  iron. 

Treatment  of  the  Precipitate.  —  At  the  cleaning  plant,  the  crude  pre- 
cipitate, containing  70  per  cent  copper,  by  means  of  a  strong  jet  of  water, 
is  gradually  worked  over  and  through  a  copper-plate  screen,  this  screen 
being  situated  at  the  head  of  a  long  launder.  The  oversize  of  the  screen, 


SHANNON  COPPER  COMPANY  PROCESS 


421 


consisting  of  leaf  copper  and  small  pieces  of  iron,  is  thrown  into  a  heap  to 
be  picked  over  by  girls  who  remove  the  scrap  iron.  The  fine  passes 
through  the  copper-plate  screen,  and  is  turned  over  by  a  stream  of  water 
that  washes  out  the  dirt  and  light  particles,  leaving  the  copper  behind. 

At  the  head  of  the  washing  launder  for  a  few  yards  is  found  No.  1 — pre- 
cipitate of  94  per  cent  Cu  and  0.3  per  cent  As.  Farther  along  is  No.  2 — 
precipitate  of  92  per  cent  Cu.  Next 
comes  No.  3 — precipitate,  which  is 
fine,  and  contains  50  per  cent  Cu, 
5  per  cent  As,  some  graphite  (from 
the  pig  iron),  and  the  bismuth  and 
antimony  precipitated  from  the 
liquor.  Nos.  1  and  2  precipitate  are 
sacked  for  shipment,  and  No.  3  is 
added  to  a  blast-furnace  matting 
charge,  the  copper  combining  to 
form  matte,  while  the  impurities 
mostly  volatilize. 

It  has  been  urged  against  the          FlQ  221.-Method  of  Removing  the 
Rio  Tinto  process  that  it  is  a  rather  Cement  Copper, 

complicated  and  very  lengthy  proc- 
ess, and  that  it  ties  up  too  much  capital.  However,  where  labor  is 
cheap,  and  forms  the  principal  expense,  where  the  ore  is  suitable  and 
abundant,  where  the  climatic  conditions  are  favorable  and  water-supply 
sufficient,  this,  as  experience  has  shown,  seems  to  be  the  most  practical  of 
the  methods  thus  far  evolved.  The  process  has  been  in  use  for  centuries. 


(2)  THE  SHANNON  COPPER  CO.  PROCESS 

This  process  is  principally  used  in  treating  oxidized  ore  together  with  sul- 
phide. A  typical  oxidized  ore  may  contain  1.9  per  cent  Cu,  40.8  per  cent 
SiC>2,  16.5  per  cent  Fe;  and  15.4  per  cent  of  the  alkali  earths.  If  sub- 
jected to  sulphuric  acid  leaching  it  would  need  7.5  to  8.8  Ib.  of  acid  per 
pound  of  copper  recovered  and  this  proved  so  expensive  that  the  present 
process  was  devised. 

In  operation,  1000  tons  of  the  oxidized  ore,  crushed  to  2-in.  size,  is  piled 
on  100  tons  of  sulphide  ore  in  circular  heaps,  and  ore  fines  are  added  as  a 
cover  to  the  thickness  of  1  ft.  There  are  ground  flues  and  stove  pipes 
placed  vertically,  reaching  from  the  sulphide  bottom  to  the  top  of  the  heaps, 
all  to  provide  draft  for  the  burning  of  the  pile.  The  sulphide  ore,  having 
been  set  on  fire,  evolves  SO2  and  SO3  gases  which  rise  through  the  entire 
pile.  At  the  same  time  the  barren  liquor  from  the  precipitation  tanks  is 
sprinkled  upon  the  heap.  Reactions  take  place  between  the  rising  gases, 


422  THE  HYDROMETALLURGY  OF  COPPER 

the  iron  sulphate  solution  and  the  oxides  and  carbonates  of  the  ore,  rep- 
resented by  the  following: 

(9)  2FeSO4+SO2  =  2Fe2(S04)3, 

and 
(10)  Fe2(S04)3+S02=2FeS04+2S03. 

That  is  to  say,  ferric  sulphate  is  alternately  reduced  to  ferrous  form  by  SO2 
and  the  resultant  ferrous  salt  again  oxidized.  In  the  reduction  sulphuric 
acid  is  set  free,  and  this  is  ready  to  dissolve  the  bases  and  the  copper  oxide. 
The  sulphur  trioxide  combines  with  the  bases  directly,  forming  sulphates. 
Toward  the  end  of  the  heap-treatment  the  pile  contains  much  ferric  sul- 
phate, an  effective  solvent  for  basic  sulphates  and  unaltered  carbonates. 
Where  the  roast  gases  most  effectively  penetrate  the  pile,  some  85  to  96 
per  cent  of  the  copper  is  brought  into  soluble  form,  but  the  considerable 
quantity  of  clayey  material  in  the  ore  causes  clogging  and  hence  imperfect 
action  in  sections  of  the  heap. 

The  ore  is  transferred  to  circular  tanks  25  ft.  diameter,  5  ft.  high  and 
holding  75  tons  each.  These  have  a  filter-bottom  covered  with  cocoa 
matting.  The  ore,  coarsely  crushed,  as  already  stated,  is  easily  leached 
with  water  by  percolation. 

The  copper  solution  from  the  leaching  tanks,  now  containing  much 
ferric  sulphate,  is  run  through  a  bed  of  raw  oxidized  ore,  followed  by  a  bed 
of  sulphide  ore  or  of  tailings  containing  sulphide.  The  action  of  the  ferric 
sulphate  on  the  oxidized  ore  is  to  dissolve  its  contained  copper,  while  the 
excess  of  ferric  salt  is  later  reduced  to  ferrous  form  as  the  result  of  its  con- 
tact with  the  sulphides. 

The  copper  liquors,  as  in  the  Rio  Tinto  process,  are  run  into  flumes  or 
launders  300  ft.  long,  5  ft.  wide,  and  2  ft.  deep,  containing  scrap  iron. 
The  spent  liquors  from  the  launders  carry  as  much  as  3.5  per  cent  Fe  as 
ferric  and  ferrous  sulphate.  The  extraction  ranges  from  73  to  82  per  cent 
of  the  contained  copper.  The  copper  precipitate  is  removed  from  the 
launders  in  the  same  way  as  in  the  Rio  Tinto  process. 

EXTRACTION  OF  COPPER  AS  A  CHLORIDE 

By  this  method  the  ore,  after  crushing  to  4-mesh  size,  is  given  a  roast 
by  which  part  of  the  sulphur  is  expelled.  At  this  stage  salt  is  added  and 
the  ore  is  finished  by  a  chloridizing  roast.  After  cooling,  the  roasted 
material  is  leached  with  a  salt  solution  to  extract  the  copper  as  chloride. 
The  copper  in  the  filtrate  is  precipitated  on  scrap  iron.  The  process  has 
the  merit,  that  where  gold  and  silver  are  present,  they  may  also  be  dis- 
solved and  recovered.  Ore  of  as  high  as  70  per  cent  silica,  containing 
sufficient  pyrite  or  chalcopyrite,  and  crushed  to  16-mesh,  can  be  given  a 


HENDERSON  PROCESS 


423 


successful  ckloridizing  roast  with  salt.  If  it  contains  no  pyrite  this  can  be 
added  and  a  chloridizing  roast  given  in  from  six  to  twelve  hours,  and  the 
copper,  gold,  and  silver  leached  out  by  means  of  water  and  dilute  acid, 
the  latter  being  obtained  as  a  by-product  of  the  roast. 

We  describe  two  methods,  viz.:  (1)  The  Henderson  process,  and  (2) 
the  Laist  process. 

1.  The  Henderson  Process. — The  well-roasted  residue,  or  cinder,  result- 
ing from  the  pyrite  used  in  making  sulphuric  acid,  contains  2  to  4  per  cent 
copper,  with  silver  and  gold.  All  these  metals  can  be  extracted  by  a 
chloridizing  roast  followed  by  leaching  with  weak  liquor  from  a  previous 
operation,  containing  water  and  dilute  hydrochloric  acid.  The  copper 
in  the  clear  filtrate  is  precipitated  upon  scrap  iron. 

The  Plant. — Fig.  232  shows  the  plan  and  a  transverse  sectional  elevation 
of  a  plant  of  the  Pennsylvania  Salt  Manufacturing  Co.,  Natrona,  Pa. 


FIG.  232.— Elevation  of  Henderson  Process  Plant. 

The  cinder  (red-roasted  or  burned  pyrite)  that  is  brought  from  the 
various  sulphuric-acid  plants  throughout  the  country  is  ground  dry  to 
20-mesh  in  a  pan-mill,  mixed  during  the  grinding  with  12  per  cent  of  the 
weight  of  salt.  This  is  raised  by  belt  elevator  to  storage  bins  on  the 
floor  K  and  put  into  the  roaster  feed  hopper  S. 

The  mixture  is  sent  to  the  five-hearth  Wedge  chloriziding  roaster, 
Fig.  233.  Each  hearth  is  perforated  with  flues  from  side  to  side  of  the 
furnace.  The  flame  from  a  firebox  on  one  side,  passing  through  these 
flues,  reaches  the  vertical  flue  on  the  opposite  side  which  leads  away  to  the 
chimney.  Dampers  regulate  the  direction  of  the  drafts.  Thus  the  com- 
bustion gases  are  kept  separate  from  the  chlorine  and  the  acid  gases  gen- 
erated from  the  ore.  Raw  pyrite  is  charged  with  the  mixture  to  have  the 
ratio  of  copper  to  sulphur  as  one  to  l£.  The  hearth  is  maintained  at  a  just 
visible  red  (525°  C.)  by  the  use  of  10  per  cent  of  fuel.  The  gas,  from  the 
ore  being  chloridized,  is  down-drafted  from  the  upper  to  the  lower  hearth, 


424 


THE  HYDROMETALLURGY  OF  COPPER 


then  passes  to  the  scrubber  or  condensing  tower  a  filled  with  lump  coke 
wet  with  a  water  spray.  The  water,  in  contact  with  the  ascending  gas, ! 
absorbs  the  chloride  and  sulphurous  acid,  reacting  as  follows : 


(10) 


C12  4-  SO2 + 2H2O  =  H2SO4 + 2HC1. 


That  is,  sulphuric  acid  and  chlorine  are  formed,  and  these  react  in  aqueous 
solution. 

The  charge  when  finished  will  contain  80  per  cent  of  its  copper  in 
soluble  form.     It  is  drawn  out  upon  the  floor,  allowed  to  cool,  shoveled 


FIG.  233. — Wedge  Chloridizing  Furnace. 

into  charge-cars,  raised  by  platform  elevator  to  the  charge-floor  level,  and 
put  into  the  leaching-tanks  d,  each  of  which  is  12  by  14  ft.  in  size. 

The  ore  is  first  lixiviated  with  a  weak  liquor  from  a  previous  operation 
to  remove  most  of  the  copper.  The  solution  becomes  a  strong  feolution. 
The  ore  is  then  treated  with  water,  to  remove  the  remaining  copper,  and 
the  solution  becomes  the  weak  solution  of  the  succeeding  operation. 
Finally,  the  weak  solution  of  hydrochloric  acid  from  the  towers  a  is  applied, 
dissolving  the  cupric  oxide  and  cuprous  chloride,  hitherto  insoluble.  The 
residue,  called  "  purple  ore,"  is  shoveled  from  the  vats  to  the  floor  c  and 
thence  discharged  into  the  railroad  cars  below. 

The  weak  solution  is  sent  to  the  lixiviation  tanks.  The  strong  solution, 
when  the  specific  gravity  reaches  18°  B.,  is  drawn  to  tanks  12  by  12  by  6  ft. 


LEACHING  PLANT  (LAIST  PROCESS) 


425 


426  THE  HYDROMETALLURGY  OF  COPPER 

filled  with  scrap  iron,  where  the  copper  is  precipitated.  The  tanks  have 
false  bottoms  of  slats  2  ft.  above  the  bottom.  Live  steam,  directed  into  the 
solution,  agitates  it.  The  copper  precipitating  upon  the  iron,  works  down 
between  the  slats  to  the  bottom  of  the  tanks  and  is  removed  to  tanks  gr, 
10  by  10  by  5  ft.  The  solution  from  this  tank  is  drawn  into  launders  con- 
taining scrap  iron  as  a  guard,  and  to  retain  any  remaining  particles  of  pre- 
cipitate. The  precipitate  is  90  per  cent  copper,  35  oz.  silver,  and  0.15 
oz.  gold  per  ton.  It  is  sold  to  the  blue  vitriol  makers,  who  pay  95  per  cent 
of  the  silver  and  the  full  value  of  the  copper  and  gold. 

The  cost  of  treatment  by  the  process,  with  common  labor  at  $1.50  per 
day,  is  $1.87  per  ton  of  cinder  treated. 

2.  The  Laist  Process. — This  process  is  used  in  the  treatment  of  mill 
tailings  containing  0.55  Cu  and  0.5  oz.  Ag  per  ton. 

The  process  consists  in  giving  these  tailings  an  oxide-chloride  roast, 
using  1  per  cent  of  common  salt.  The  roasted  ore  is  then  leached,  first 
with  a  No.  1  or  weak  solution  containing  3.5  per  cent  H2SO4  and  10  per 
cent  salt,  then  with  a  No.  2  or  strong  solution  of  6  per  cent  H^SCU  and  10 
per  cent  salt.  The  copper,  dissolved  by  the  weak  solution,  goes  to  tanks 
for  precipitation  on  scrap  iron;  the  strong  solution,  after  use,  is  returned 
to  be  employed  as  No.  1  or  weak  solution  on  the  next  charge. 

In  Figs.  234  and  235  are  given  views  of  the  leaching  plant  for  the 
treatment  of  60  tons  of  tailing  daily. 

Three  bins,  one  A  for  salt,,  one  B,  for  coal,  and  a  third  C,  having  a 
hopper  bottom,  are  for  the  storage  of  the  mixture  of  sand  and  slime  tailing 
(4  of  sand  to  1  of  slime)  which  is  to  be  treated.  These  tailings  contain 
0.6  per  cent  Cu,  82.2  per  cent  Si02, 1.9  per  cent  Fe,  2.2.  per  cent  S  and  carry 
0.55  oz.  Ag  and  0.002  oz.  Au  per  ton. 

From  the  sand  bin  the  material  is  delivered  by  a  belt-feeder  to  a  ver- 
tical elevator  discharging  into  the  feed-hopper  of  the  20-ft.  MacDougall 
roasting  furnace  D.  Referring  to  Fig.  236,  the  furnace  is  of  the  six-hearth 
type  with  a  lower  water- jacketed  one  for  cooling  the  calcine  or  roasted  ore. 
There  are  two  fireboxes  with  shaking  grates,  discharging  into  hearth  No.  2, 
which  thus  becomes  a  combustion  chamber,  where  the  coal  gases  burn  with 
along  flame  and  the  products  of  combustion  are  drawn  off  by  a  No.  11 
Buffalo  blower  to  the  chimney.  On  the  upper  three  hearths  or  floors  the 
tailings  are  roasted  and  brought  to  the  temperature  of  540°  C.  falling  then 
to  the  floor,  No.  4  where  1  per  cent  of  common  salt  is  fed  in.  During  their 
passage  over  the  fourth,  fifth,  and  sixth  floors  the  copper,  as  well  as  the  sil- 
ver compounds,  are  chloridized,  the  heat  still  present  in  the  ore  being  suf- 
ficient to  ensure  the  chloridizing  reactions.  A  small  volume  of  air  is 
drawn  through  hearths  Nos.  6  and  7  by  a  No.  1O  Buffalo  blower  to  an 
absorption  tower,  E,  so  as  to  catch  any  copper  or  silver  that  has  been  vola- 
tilized. This  tower,  filled  with  coke,  is  showered  with  water  delivered  to 


LEACHING  PLANT  (LAIST  PROCESS) 


427 


i 


428 


THE  HYDROMETALLURGY  OF  COPPER 


it  by  a  IJ-in.  bronze  pump.     To  withstand  the  chlorine  fumes  evolved 

from  the  cooling  ore  on  the  seventh  hearth,  copper  rabbles  are  provided. 

The  cooled  ore  is  delivered  by  a  horizontal  screw-conveyor  to  a  ver- 


42  Flue  holeg  2*r  6*in 
Circ.  unequally  spaced 


Siiisiyil 

FIG.  236. — Sectional  Elevation  of  Roasting  Furnace. 

tical  elevator  which  discharges  it  upon  a  12-in.  belt-conveyor  to  the  sand 
distributor  of  one  of  the  32  by  12  ft.  "  leaching  tanks."  The  distributor, 
revolving  about  the  axis  of  the  tank,  delivers  the  calcine  evenly  over  its 
whole  area.  When  ore  has  been  roasted  and  its  colloids  destroyed,  it 


THE  LA1ST  PROCESS  429 

is  in  excellent  condition  for  leaching.  The  leaching  tanks  will  hold  350 
to  400  tons. 

There  are  two  lead-lined  solution-tanks  each  27  ft.  diameter  by  12  ft. 
deep.  "  No.  1  solution  tank  "  contains  the  No.  1  or  weak  solution  of 
3J  per  cent  SO4  and  10  per  cent  of  common  salt;  "No.  2  solution  tank  " 
carries  the  strong  solution  of  6  per  cent  H2SO4  and  10  per  cent  salt.  For  a 
320-ton  charge,  the  No.  1  solution  is  run  on  the  tailing  first  to  get  out  the 
bulk  of  the  copper,  and  remains  in  contact  with  it  for  fourteen  hours. 
This  solution,  the  only  one  precipitated,  is  run  out  by  a  launder  to  one  of 
the  6-in.  Pohle  air-lifts  Py  P,  which  raises  it  so  that  it  is  carried  to  the  27 
by  12-ft.  "  copper  solution  tank."  From  this  tank  it  flows  in  a  regulated 
stream  to  one  of  the  "  lead-lined  precipitating  launders  "  at  the  end  of  the 
building  filled  with  scrap  iron,  where  the  copper  and  silver  are  quite  pre- 
cipitated. The  spent  or  barren  liquor  flows  to  waste.  The  precipitate, 
when  a  clean-up  is  made,  is  washed  down  into  the  "  clean-up  tank,"  and 
the  precipitate  collected  for  further  treatment. 

To  return  to  either  leaching  tank:-  Twenty-four  tons  of  strong  or  No.  2 
solution  is  run  on,  and  stands  for  seventy-two  hours,  after  which  it  is 
returned  to  "  No.  1  solution  tank  "  as  weak  or  No.  1  solution.  Following 
this  the  remaining  values  are  removed  with  water-washes.  The  final 
tailings  are  sluiced  out  of  the  tank  with  a  2-in.  hose,  and  sent  to  waste  by 
means  of  a  launder  carrying  a  3-in.  stream  of  water.  The  arrangement 
of  the  launders  and  of  the  four  discharge  valves  of  a  tank  are  well  shown  on 
the  plan. 

The  percentage  of  recoverable  copper  is  85.4  and  of  silver  91.1. 

SULPHURIC  ACID  LEACHING 

This  method  of  extraction  is  suited  to  a  limited  range  of  ores,  those 
which  will  not  consume  much  acid  (generally  lost  in  uselessly  dissolving 
bases)  and  those  in  which  the  copper  minerals  are  present  in  soluble  form. 
As  regards  the  first  objection,  there  are  oxidized  copper  ores  containing 
iron  oxide  and  especially  the  alkali  earths  which  consume  acid.  The 
copper  minerals  unattacked  by  sulphuric  acid  are  metallic  copper,  cuprite, 
fresh  unaltered  sulphides  such  as  chalcopyrite  and  covellite,  and  massive 
chrysacolla.  When,  however,  the  ore  contains  malachite,  azurite,  copper 
oxide,  and  basic  sulphates  mainly,  then  it  may  be  quite  suited  to  sulphuric- 
acid  treatment. 

Although  we  have  a  suitable  ore,  still  the  acid  will  act  on  the  clayey 
minerals,  the  iron  oxides  and  the  alkaline  earths  so  that  these  bases  should 
be  present  in  small  quantity  only.  The  waste  of  acid  is  not  the  only  draw- 
back; the  bases  named  accumulate  in  the  solution,  which  naturally  is  to  be 
re-used,  and  finally  the  whole  has  to  be  run  to  waste  and  fresh  acid  used. 


430  THE  HYDROMETALLURGY  OF  COPPER 

When,  to  save  scrap  iron,  electrolytic  precipitation  is  employed  copper 
is  indeed  precipitated,  but  the  SOs,  released  from  the  electrolyte,  oxidizes 
the  ferrous  oxide  present  to  ferric  form  and  this  at  once  proceeds  to 
redissolve  the  copper.  True,  porous  diaphragms  have  been  used  to  confine 
the  ferric  oxide  to  the  anodes,  but  this,  aside  from  added  expense,  increases 
the  resistances.  Farther,  no  anode  is  altogether  satisfactory.  If  of  lead, 
this  is  gradually  changed  to  peroxide  by  the  oxidizing  effect  mentioned, 
and,  while  the  peroxide  may  be  recovered  and  again  reduced  to  lead,  this 
increases  working  costs. 

The  electrolyte  becomes  foul,  owing  to  the  accumulation  of  sulphates 
of  iron  and  other  metals,  and  it  is  necessary  periodically  to  send  some  of  the 
solution  to  waste,  thus  causing  a  loss  of  acid.  Regeneration  of  the  acid 
is  effected,  and  in  fact,  when  sulphur  dioxide  is  injected  into  the  electrolyte 
as  a  depolarizer,  an  excess  of  acid  is  obtained  by  the  combination  of  the  SC>2 
with  the  nascent  oxygen  liberated  at  the  anode;  still  so  many  reactions 
occur  among  the  foreign  metals  present  in  the  electrolyte,  and  so  much 
trouble  has  been  found  in  properly  regulating  the  current  density,  that  a 
great  deal  of  current  is  wasted  in  excess  of  that  theoretically  needed.  In 
the  deposition  from  a  copper  sulphate  solution  this  would  be  2.14  Ib.  of 
copper  per  kilowatt-hour,  but  in  practice  but  50  per  cent  of  this  has  been 
obtained.  With  a  pure  electrolyte  and  taking  proper  precautions  to  pre- 
vent wastage  at  the  cathode  an  efficiency  of  90  per  cent  or  2  Ib.  Cu  per 
kilowatt-hour  should  be  attained.  These  precautions  would  consist  in 
precipitating  the  interfering  metals  by  chemical  means  before  the  solution 
goes  to  the  electrolytic  cells.  With  such  efficiency  the  cost,  especially 
where  hydro-electrolytic  power  is  available,  should  not  exceed  1  cent  per 
pound  of  copper  deposited.  The  advantages  of  electrolytic  precipitation 
are  that  the  acid  is  regenerated  and  that  a  pure  copper  is  produced, 

We  describe  herewith  two  methods  which  have  worked  well  upon  suit- 
able ores: 

(1)  The  Butte-Duluth  process. 

(2)  The  Ajo  Process. 

(1)     THE  BUTTE-DULUTH  PROCESS 

The  ore  exists  as  a  large  body  of  decomposed  granite  carrying  2  per 
cent  of  copper  as  copper  carbonates  chrysacolla  and  cuprite. 

Briefly  stated:  The  crushed  ore  is  leached  with  sulphuric  acid,  the 
filtrate  heated  to  60°  C.,  the  copper  electrolytically  deposited,  and  the 
remaining  solution,  still  containing  copper,  strengthened  with  acid  and 
returned  for  re-use.  The  ore  is  farther  leached  with  water,  the  water  run 
through  other  ore,  then  over  scrap  iron  to  obtain  the  remaining  copper. 

The  ore,  crushed  to  suitable  size  for  leaching  as  described  under  crush- 


BUTTE-DULUTH  PROCESS 


431 


ing,  is  fed  to  three  rectangular  leaching  tanks,  70  ft.  long,  12  ft.  wide  by 
6  ft.  deep,  lined  with  sheet-lead  and  having  plugged  bottom-openings  12 
in.  diameter.  There  is  a  false  or  filter  bottom  made  of  2-in.  planks,  bored 
full  of  f-in.  holes,  for  the  passage  of  the  solution.  The  leaching  tanks, 
when  filled,  are  treated  with  a  10  per  cent  solution  of  sulphuric  acid, 
which  remains  upon  the  ore  for  twenty-four  hours,  dissolving  most  of  the 
copper.  This  copper-bearing  solution,  still  containing  5  to  8  per  cent 
H2SO4,  and  having  2  per  cent  Cu,  is  drawn  from  the  tanks  and  passes  to  a 
storage  sump.  From  the  storage  sump  it  is  lifted  by  steam-lifts  into  the 


I    Storage  Tank    [      j" 


I    Water  Tank  ] 


Jill 


!• 

lutti 
CY 

11, 

c 

1    1  \. 

Sulphuric  Acid  for 
'  rdiiation     | 


r   Sulphur! 
S"°d' 


Sump  Tacks 


LEGEND 

—  Mill-Solution 

—  Wash  Waters 
=  Sulphuric  Acid 


FIG.  237. — Solution  Flow-sheet,  Butte-Duluth  Mill. 

temperature  cells,  these  heating  the  solution  to  the  desired  temperature  of 
60°  C.     The  temperature  cells  are  lead-lined  wooden  tanks. 

From  the  temperature  cells  the  solution  flows  through  the  electrolytic 
cells,  where  part  of  the  copper  is  deposited  and  acid  regenerated.  After 
flowing  through  these  cells,  the  H2S04  has  been  raised  to  about  1  per  cent 
in  strength.  The  cells  (shown  as  a  rectangle  divided  into  four)  consist  of 
twelve  cells  each  8  ft.  long  30  in.  wide  and  39  in.  deep,  lined  with  4-lb.  hard 
lead.  In  each  cell  are  twenty  anodes  of  hard  lead  weighing  10  Ib.  per  square 
foot  and  nineteen  cathodes,  being  starting  sheets  weighing  2  to  3  Ib.  when 
first  placed.  These  remain  in  the  cells  seven  to  ten  days.  They  grow  to 
40  to  60  Ib.  before  removal  and  assay  99.96  per  cent  Cu.  The  anodes  and 
cathodes  are  electrically  connected  hi  multiple,  the  cells  in  series. 


432  THE  HYDROMETALLURGY  OF  COPPER 

The  solution  passes  to  the  sump  tanks  where  it  is  strengthened  to  10 
per  cent  H2SO4,  the  acid  being  received  from  a  tank  marked  "  sulphuric 
acid  for  standardization."  From  the  sump  tanks  the  strengthened  solu- 
tion is  pumped  to  the  storage  tank  placed  40  ft.  higher,  so  as  to  connect  the 
"  leaching  tanks  "  by  gravity.  The  pumping  is  effected  by  the  aid  of  two 
4-in.  lead-lined  centrifugal  pumps. 

Returning  to  the  leaching  tanks :  The  ore  has  just  been  leached  with  the 
10  per  cent  solution  coming  from  the  storage  tank  above.  This  is  followed 
by  several  water-washes  from  the  "  water  tank."  The  first  wash-water 
containing  sulphuric  acid  is  added  to  the  mill  solution  of  the  storage  sump. 
The  remaining  washes,  weak  in  acid,  go  to  the  leaching  tank  shown  at  the 
right  side  of  the  flow-sheet.  In  order  to  keep  the  mill  solution  pure,  and 
prevent  it  from  accumulating,  a  quantity  of  it,  equal  to  the  first  wash-water, 
is  passed  by  the  line  marked  "  a  small  part  sol."  through  ore  in  the  right- 
hand  leaching  tank  until  its  contained  acid  is  used  up  by  acting  on  the  ore, 
then  run  to  the  launder  marked  "  scrap-iron  precipitation";  the  spent 
solution  is  wasted.  Thus  part  of  the  copper  is  recovered  electrolytically 
and  a  small  part  by  means  of  scrap  iron. 

The  costs  are  thus  roughly  given  per  pound  of  electrolytic  copper  pro- 
duced on  the  treatment  of  50  tons,  recovering  2000  Ib.  of  copper  daily: 

3£  Ib.  acid  at  $27  per  ton • $0 . 04725 

Power  for  crushing  and  electrolytic  deposition 0.01 

Management  and  labor  $160  daily 0 . 08 


$0.13725 

The  cost  for  acid  should  be  greatly  reduced,  and  the  labor  costs  are  exces- 
sive, due  to  construction  work  and  alterations  in  addition  to  actual 
operating. 

(2)     THE  AJO  PROCESS 

The  Process. — This,  in  brief,  consists  in  leaching  the  ore  (crushed  to 
J-in.  size)  for  eight  days  by  a  counter-current  system;  reducing  the  ferric 
iron  in  the  resultant  solution  to  ferrous  form,  using  sulphurous  acid  gas 
(862)  to  do  so;  and  electrolytically  precipitating  out  part  of  the  copper, 
which  is  then  returned  to  the  leaching  solution. 

Coarse  Crushing. — This  is  done  during  two  eight-hour  shifts,  since 
there  is  no  storage  between  the  plant  and  the  mine.  The  ore,  some  of 
3  to  4  ft.  in  minimum  dimension,  is  crushed  in  two  sets  of  gyratory  ore- 
breakers  to  4-in.  size,  then  delivered  to  a  storage-bin  of  10,000  tons  aggre- 
gate capacity,  as  is  clearly  shown  by  the  general  flow-sheet,  Fig.  238. 

Fine  Crushing. — This  is  done  between  3  P.M.  and  7  A.M.,  but  can  be 
kept  up  for  twenty-four  hours  if  necessary.  The  ore  is  crushed  in  two 


THE  AJO  PROCESS 


433 


Oement  Copper 

for  Shipment 

63.54  On. 


Tank  House  Return 


LEGEND 

Pj-  Circulating  and 

Advance  Pumps. 
P3-  Make-np  Advance  Pump 
P4-  Spray  Advance  Pnmp. 
P5-  Settling  Tank  Pump. 
P6.  Pnmp  to  Tank  House. 
P7-  Tank  House  Return  Pump-. 
Si-  S2  &  S3-  SO2  Towers. 

Wi   Wash  Water  No.l. 

W2  Wash  Water  No.2. 

W3  Wash  Water  No.3. 

W4  Wash  Water  No.4. 

f  Return  from  Tank  House. 
— O— (  Make-up  Solution,  and 

(Advance  from  Tank  to  Tank. 
•       Circulation  on  Leaching. 
Tanks. 

— Advance  to  Tank  House. 

•ooooSOs  Gas. 

Arrows  Indicate  Direction  of  Flow. 


2  I  • 


434  THE  HYDROMETALLURGY  OF  COPPER 

stages  by  means  of  Symond's  disk  crushers,  the  ore  being  drawn  off  and 
crushed  according  to  the  immediate  needs  of  the  leaching  plant. 

Leaching. — Of  the  twelve  tanks  eleven  are  for  leaching,  the  twelfth  is  a 
solution  settler.  Of  the  eleven,  seven  always  contain  ore  in  process  of 
leaching.  Referring  to  the  flow-wheel,  we  may  assume  that  the  ore  in 
tank  No.  10  is  the  oldest,  and  No.  5  the  newest  in  the  circuit;  the  No.  6 
is  being  charged  with  ore,  No.  7  is  empty,  No.  8  is  being  excavated  and  No.  9 
in  various  stages  of  washing  and  draining.  When  tank  No.  6  has  been 
charged,  and  it  is  ready  for  the  leaching  cycle,  the  "  acid  advance,"  that 
is,  the  amount  of  acid-bearing  solution  that  proceeds  from  tank  to  tank  is 
increased  to  its  maximum  amount  of  2000  gal.  per  minute  for  four  hours, 
this  solution  being  gotten  from  storage  tank  A  or  E.  Meanwhile  the  usual 
advance  of  1000  gal.  per  minute  continues  to  go  from  tank  No.  5  to  six 
reducing  towers  marked  in  pairs  Si  to  83,  where  it  is  subject  to  the  reducing 
action  of  SO2  in  water  solution.  The  excess  of  1000  gallons  is  advanced 
into  tank  No.  6  until  the  ore  is  covered  with  it — this,  in  order  to  prevent 
any  interruption  of  flow  to  the  towers.  When  the  ore  is  covered  the  excess 
advance  is  cut  off  to  the  normal  of  1000  gal.  per  hour.  Solution  on  the  new 
charge  is  now  circulated  on  itself,  until  it  is  clarified,  or  for  about  four 
hours.  Tank  No.  6  is  now  put  in  circuit  and  the  neutral  advance  (acid 
free  solution)  to  the  tower  comes  off  from  tank  No.  6  in  place  of  tank  No.  5. 

The  leaching  of  the  ore  in  tank  No.  6,  now  begun,  continues  for  seven 
days,  during  which  the  free  acid  in  the  solution  increases  from  0.5  per  cent 
to  3.0  per  cent  on  the  seventh  day.  At  the  end  of  the  seventh  day,  the 
"  acid  advance  "  from  the  tank  house  is  transferred  from  tank  No.  10  to 
tank  No.  11.  Upon  the  entrance  of  a  new  charge  into  the  circuit,  the  solu- 
tion remaining  in  the  oldest  tank  is  drained  to  solution  storage,  where  it  is 
standardized  by  additions  of  sulphuric  acid  and  is  later  used  as  "  acid 
advance."  After  draining  the  tank  is  ready  for  wash- water.  As  the 
copper  that  is  taken  away  in  the  leaching  is  about  two-thirds  the  total, 
the  question  of  thorough  washing,  to  remove  the  rest,  is  important.  Four 
successive  washings  with  the  drainings  between  are  used.  During  the 
three-hour  circulation  that  each  wash  is  given,  an  equilibrium  between  the 
dissolved  copper  in  the  tailings  and  that  of  the  wash-water  being  applied, 
is  expected  to  be  reached.  To  follow  more  readily  the  method  of  washing 
a  charge  the  flow-sheet  must  be  referred  to.  When  tank  No.  9  has  been 
thoroughly  drained,  the  charge  is  covered  with  wash-water  from  W, 
circulated,  then  drained  to  solution  storage  tank  A  or  E;  this  constitutes 
the  first  wash.  It  is  now  covered  with  wash-water  from  W2,  similarly 
circulated  and  drained  to  W.  In  the  same  way  the  wash-water  from  Wz  is 
put  on,  circulated  and  drained.  The  fourth  or  last  wash,  consisting  en- 
tirely of  fresh  water,  is  pumped,  circulated  and  drained  into  wash-water 
tank  W%.  In  this  manner  the  fourth  wash  of  any  one  charge  is  used  as 


THE  AJO  PROCESS  435 

the  third  wash  of  the  succeeding  charge,  the  third  as  the  second,  and  the 
second  wash  as  the  first.  In  other  words  each  wash-water  is  used  four  times, 
the  copper  contents  increasing  each  time,  when  it  enters  the  system  and 
makes  up  for  the  continuous  losses  of  solution  due  to  evaporation,  to  dis- 
card and  to  about  1 1  per  cent  of  solution  taken  away  in  the  tailings.  Thus, 
before  the  first  wash  the  solution  contained  2.56  per  cent  acid  and  2.4  per 
cent  copper,  while  at  the  end  of  the  fourth  wash  there  remained  0.10  per 
cent  acid  and  0.38  copper.  To  obtain  an  even  better  extraction  of  the 
copper  it  has  been  proposed  to  give  a  fifth  wash,  then  allowing  the  result- 
ant solution  to  flow  over  scrap  iron. 

Arrangement  of  the  Leaching  System. — The  twelve  leaching  tanks 
are  arranged  in  two  rows,  as  shown  on  the  flow-sheet,  and  more  particu- 
larly as  seen  in  the  sectional  transverse  section,  Fig.  239 .  The  aisle  between 
the  two  rows  of  tanks  is  108  ft.  wide  and  contains  what  is  called  the  "  cen- 
tral structure  "  of  the  same  length  as  the  row  of  tanks.  This  consists  of 
six  heavy  concrete  piers  each  of  four  pillars  supporting  steel  trusses  from 


ROBIN'S  SPREADING  BRIDGE  s^Sc^uRE  ""LETT   EXCAVATOR 


FIG.  239. — Cross-section  of  Leaching-tanks. 

pier  to  pier.  The  structure  has  two  decks,  the  upper  carrying  the  belt- 
conveyor,  the  lower  the  solution  launders  and  the  pipe-lines.  At  each 
concrete  pier  are  four  pumps  and  pipe  connections.  Underneath  the 
central  structure  and  parallel  to  it  are  to  be  seen  the  two  drainage  launders 
used  in  carrying  the  solutions  from  the  leaching  tanks  to  the  solution 
storage. 

The  ore,  from  storage  as  finally  crushed,  is  conveyed  through  an  auto- 
matic sampling  plant,  and  thence  by  a  main  conveying-belt  the  full  length 
of  the  central  structure.  To  fill  any  desired  tank  there  is  a  tripper  on  this 
main  conveyor  that  delivers  to  the  belt  of  a  Robins  spreading  bridge. 
This  bridge  set  over  any  desired  tank  delivers  its  load  into  it  from  side  to 
side  and  can  be  moved  to  fill  any  part  of  the  tank. 

Removal  of  Tailings. — After  a  charge  has  been  washed  and  drained, 
the  tailings  are  removed  by  a  Hulett  excavator,  similar  to  those  used  in 
unloading  iron  ore  at  the  lower  lake  ports.  A  heavy  steel  bridge  on  tracks 
spans  the  leaching  tanks  and  can  travel  their  entire  length.  On  this  bridge 
travels  the  excavator,  consisting  of  a  walking-beam,  bucket-leg,  and  bucket 


436  THE  HYDROMETALLURGY  OF  COPPER 

of  12  tons  capacity.  This  will  unload  the  tank  at  the  rate  of  500  tons  per 
hour.  Two  eight-car  trains  are  released  from  mine-service  at  4  P.M.  for 
the  transportation  of  tailings,  and  twenty-one  to  twenty-three  train  loads 
are  required  when  removing  the  contents  of  a  tank.  When  it  is  desired  to 
exchange  places  for  the  spreading  and  excavator  bridges  it  is  thus  per- 
formed: Just  beyond  the  last  tank  is  a  transfer-table  track  and  just 
beyond  the  transfer-pit  are  tail-tracks  matching  the  bridge  tracks.  The 
unloading  bridge,  for  example,  is  run  over  the  transfer  table  upon  its  tail- 
track.  The  transfer  table  is  then  moved  to  match  the  excavator  bridge, 
which  is  then  run  upon  it,  moved  over  to  the  other  set  of  tracks  and  set  in 
place  ready  for  unloading  a  tank.  The  unloading  bridge  in  its  turn  is  put 
on  the  transfer  table,  transferred  and  set  in  place  on  the  other  set  of  tracks. 
SO2  Towers. — In  the  electro-deposition  of  copper  from  a  sulphuric  acid 
solution  any  iron  in  ferric  form  will  be  reduced  by  the  current  to  ferrous 
form,  thus  using  up  electric  energy.  To  overcome  this  SO2  gas  was 
employed  to  bring  the  ferric  iron  into  ferrous  form  according  to  the  equa- 
tion 

(11)  Fe2O3+S02+H2O  =  2FeO+H2SO3. 

Where  care  was  taken  to  send  to  the  862  towers  neutral  or  slightly  acid 
solutions,  this  proved  easy. 

Roasting  for  SC>2. — Referring  again  to  the  flow-sheet,  Fig.  238,  there  are 
six  seven-hearth  roasters  which  carefully  roast  75  tons  daily  of  Bisbee 
pyrite  ore.  The  strong  gas,  containing  8  to  10  per  cent  862,  leaving  the 
roasters  passes  to  a  Cottrell  precipitator  or  treater,  where  it  is  cleaned 
from  dust  before  it  enters  the  spray  or  cooling  chamber.  Upon  the  top 
and  sides  of  the  chamber  are  nozzles  by  which  about  100  gal.  per  minute  of 
"  neutral  advance  "  solution  is  sprayed  to  cool  the  gas  before  it  enters  the 
reducing  towers.  The  ferric  iron  in  this  solution  is  at  once  reduced  to 
ferrous  form. 

There  are  six  towers  built  of  sheet-leaci  and  arranged  in  pairs.  They 
are  20  to  28  ft.  diameter  by  40  ft.  high  and  rest  on  a  concrete  base,  the 
lower  edges  dipping  into  a  lead-lined  sump,  6  ft.  deep,  so  that  the  edge 
below  the  surface  of  the  liquid  in  the  pan  forms  a  seal  against  the  escape  of 
the  gas.  The  tower  space  is  filled  with  boards  cross-piled  on  edge  and  1  in. 
apart.  The  solution  comes  to  the  top  of  the  tower  by  launder  and  enters 
it  through  gas  seals,  that  is  a  siphon-shaped  pipe,  that  lets  the  solution 
through  without  letting  the  gas  escape.  The  solution  going  through 
these  seals  trickles  down  through  the  boards,  wetting  them,  and  is  subject 
to  intimate  contact  with  the  strong  gas,  cooling  it  to  atmospheric  tempera- 
ture, and  at  the  same  time  reducing  all  the  iron  present  to  ferrous  form. 
Between  the  second  and  third  towers  is  a  60-in.  suction-fan  made  of  lead. 
This  draws  the  gas  from  the  roasters  through  the  Cottrell  treater,  spray- 


THE  AJO  PROCESS  437 

chamber,  and  third  set  of  towers,  and  forces  it  through  the  second  and  first 
sets  to  the  atmosphere. 

The  solution  or  neutral  advance,  say  900  gal.  per  minute,  travels  counter- 
current  to  the  flow  of  gas,  that  is,  the  most  reduced  solution  comes  in  con- 
tact with  the  strongest  gas.  The  solution  from  the  newest  tank  of  ore  is 
pumped  to  the  top  of  the  third  pair  of  towers  (83,  83)  by  a  9-in.  centrifugal 
pump.  From  the  bottom  of  these  it  is  lifted  to  the  top  of  82,  82,  and  from 
their  pumps  to  the  top  of  8,  8.  From  the  bottom  of  these  it  is  pumped  by 
PS  to  the  settling  tank,  the  fourth  tank  of  the  nearest  row.  The  purpose 
of  this  tank  is  two-fold;  to  settle  out  the  slime  and  to  cause  additional 
reduction  as  the  solution  stands.  The  average  of  ferric  iron  in  the 
solution  entering  the  towers  is  0.80  per  cent  and  of  that  leaving  the  set- 
tling tank  to  go  to  the  electrolytic  tank  house  only  0.10  per  cent  Fe2Oa. 

Electrolytic  Deposition. — The  electrolytic  tank  house  is  arranged  much 
as  in  electrolytic  copper  refining,  which  see.  The  tanks  are  arranged  in 
banks  with  sills  between.  There  are  12  banks  of  10  tanks  each  and  4 
banks  of  8  tanks  each.  Each  tank  has  84  anodes  and  77  cathodes  per  tank. 
The  anodes  are  of  hard  lead,  containing  3.5  per  cent  antimony;  the 
cathodes  are  copper  starting-sheets,  originally  15  to  18  Ib.  in  weight,  while 
the  finished  cathodes  weigh  130  to  140  Ib.  each.  Of  the  total  of  152  tanks, 
25  are  reserved,  being  employed  making  starting-sheets,  the  remaining 
127  tanks  are  depositing  copper  on  the  sheets  for  the  production  of  cathodes. 
The  starting  blanks  upon  which  the  starting  sheets  are  made  are  of  anti- 
monial  or  hard  lead  like  the  anodes. 

The  electric  current  is  passed  through  each  tank  in  parallel  and  through 
them  all  in  series.  It  is  supplied  by  two  identical  units  each  of  15,000 
amperes  and  each  having  seventy-six  tanks  in  series. 

The  Electrolyte. — The  solution  entering  the  tank-house  contains  Cu 
3.0  per  cent,  Fe  in  ferrous  form  2-3  per  cent,  Fe  in  ferric  form  0.085  per 
cent.  When  it  leaves  the  tank  house* to  go  to  T.  H.  R.  (tank  house  return) 
sump  its  copper  has  fallen  to  2.5  per  cent,  that  is,  0.50  per  cent  has  been 
deposited,  the  ferrous  iron  has  decreased  to  1.66  per  cent  and  the  ferric 
iron  correspondingly  increased  to  0.75  per  cent.  Sulphuric  acid  has  also 
been  set  free,  there  being  2.10  per  cent  as  against  1.70  per  cent  entering  the 
tank  house.  The  average  weight  of  copper  deposited  per  kw.  hour  may 
be  given  at  0.8  Ib. 

As  will  be  seen  in  the  flow-sheet  the  tank-house  return-solution  is  made 
up  by  addition  of  solution  from  tanks  E  and  A  and,  if  necessary,  the  acid 
in  it  is  strengthened  by  the  addition  of  strong  acid  made  at  another  plant. 
This  is  in  order  that  the  strongest  acid  solution  shall  go  to  the  newest  tank. 
The  new  acid,  that  brought  from  the  outside,  will  amount  to  60  per  cent 
of  the  total  acid  consumed.  Of  the  remaining  40  per  cent  some  32  per  cent 
is  regenerated  in  the  SO2  towers  and  8  per  cent  at  the  electrolytic  tanks. 


438  THE  HYDROMETALLURGY  OF  COPPER 

Only  about  half  of  the  total  acid  used  in  an  eight-day  leaching  is  utilized 
in  dissolving  copper,  the  remainder  is  used  up  in  dissolving  impurities. 
Therefore  these  impurities  gradually  accumulate,  making  the  solution 
more  and  more  sluggish.  To  keep  the  solution  active  a  portion  must  be 
discarded  and  replaced  by  fresh  water  introduced  at  the  last  wash.  This 
daily  discard  must  be  such  that  its  impurities  shall  be  equivalent  to  those 
taken  up  each  day.  In  this  particular  plant  it  amounts  to  90  gal.  per  min- 
ute out  of  the  1324  of  neutral  advance. 

The  copper  in  the  discarded  solution  is  recovered  by  passing  through 
scrap-iron  boxes  or  launders,  just  as  described  for  the  Rio  Tin  to  process, 
which  see.  It  takes  2  Ib.  of  iron  scrap  to  precipitate  one  of  copper.  The 
resultant  precipitate,  which  contains  perhaps  75  per  cent  copper,  is  shipped 
away  for  smelting.  About  10  to  15  tons  of  copper  is  recovered  in  this  way 
daily.  The  cathodes  from  the  electrolytic  tank  house  are  quite  pure,  con- 
taining as  high  as  99.85  per  cent  copper. 

AMMONIA  LEACHING 

Limitations  of  the  Process.  —  This  is  suited  to  the  treatment  of  a 
mixture  of  sulphide  and  oxidized  ores,  that  may  contain  so  much  calcite 
and  dolomite  as  to  forbid  acid  leaching.  The  sulphides  are  removed  by 
concentration  and  the  tailings,  containing  the  copper  carbonates  and  oxides, 
are  leached  out  with  a  carbonate  of  ammonia  solution.  Where,  as  in  the 
Lake  Superior  region,  the  copper  occurs  native,  then  this  is  likewise  re- 
moved by  concentration  except  the  finest  particles  of  native  copper  which, 
with  any  copper  oxide,  may  be  dissolved  by  the  ammonium  carbonate. 
In  the  case  of  malachite  we  have  the  reaction 


(12)  CuCO2Cu(OH2)+2(NH4)2CO3  =  CuCO3^  2NH3+H2O+2CO2, 

the  CO2  gas  escaping  freely.  The  reaction  with  azurite  is  similar.  The 
fine  copper  particles  with  free  access  of  air  are  oxidized  to  copper  oxide  and 

(13)  CuO+(NH4)2CO3  =  CuCO3,    2NH3+H2O. 

Where  the  air  does  not  have  free  access,  the  cupric  salt  thus  formed  comes 
in  contact  with  the  oxidized  copper  particles  and  is  reduced  to  the  cuprous 
state. 

(14)  CuCO3+Cu  =  Cu2CO3,  2NH3. 

However,  this  latter  compound,  coming  where  there  is  access  of  air,  again 
rapidly  oxidizes  to  cupric  form,  and  can  then  proceed  to  dissolve  an  addi- 
tional quantity  of  native  copper. 

Upon  subjecting  the  solution  to  boiling  by  steam,  the  ammonia  and 
carbon  dioxide  of  the  cupric  carbonate  are  distilled  off  thus  : 

(15)  CuC03,  2NH+H20  =  H2O+heat+CuO+  (NH 


AMMONIA  LEACHING  439 

The  copper  oxide  falls  out  of  the  solution  as  a  heavy  powder,  and  the 
ammonium  carbonate  is  absorbed  by  water  and  this  recovered  for  further 
use. 

AMMONIA  LEACHING  AT  KENNICOTT 

We  present  in  Fig.  240  the  plans  and  elevations  of  an  ammonia-leaching 
plant  for  the  treatment  of  tailings  containing  1.46  per  cent  copper  or 
of  1.14  per  cent  copper  in  soluble  form  as  copper  carbonates.  After 
ammonia  treatment  the  residue,  going  to  waste,  still  retained  0.26  per 
cent  of  copper  indicating  a  recovery  of  77  per  cent  of  the  soluble  portion. 

Referring  to  Fig.  240,  section  Z,  Z,  the  tailings  are  brought  by  a  launder 
to  the  275-ton  storage  bin.  Thence,  when  needed  to  fill  the  leaching  tanks 
(see  the  elevation),  they  are  raised  by  a  vertical  belt  elevator,  and  after  a 
Vezin  sampler  has  taken  out  a  portion,  are  transferred  by  a  short,  inclined 
conveyor  to  the  long  distributing  conveyor,  seen  in  the  plan  running 
over  the  leaching  tanks.  When  leached,  the  exhausted  tailings  are 
removed  by  a  tailings-discharge  conveyor  beneath.  The  tanks  are  30 
ft.  diameter  and  will  hold  500  tons.  They  are  filled  from  the  distributing 
conveyor  by  a  revolving  plate  provided  to  distribute  the  feed  evenly  over 
the  whole  area  and  to  give  a  uniform  mixture  of  coarse  and  fine.  A  dome- 
shaped  cover  is  then  put  on. 

The  first  leaching  solution,  consisting  of  rich  copper-ammonia  solution 
from  a  previous  leach,  together  with  concentrated  ammonia  from  the  still  is 
run  into  the  tank  at  the  bottom,  and  rises  through  the  charge,  displacing 
the  moisture  of  the  tailings  before  it.  The  first  part  of  this  moisture, 
practically  barren,  is  sent  to  waste.  Rich  solution  is  now  brought  on  top 
of  the  charge  and  leaches  downward.  As  fast  as  the  downward  flow  con- 
tinues it  is  returned  fron  below  by  means  of  a  centrifugal  pump  to  the  top 
of  the  charge,  this  being  kept  up  as  long  as  the  solution  will  dissolve  the 
copper,  or  for  about  thirty  hours.  The  leaching  solution,  carrying  4J 
per  cent  Cu  and  7.5  per  cent  NHa,  is  now  drawn  off,  part  going  to  the  evap- 
orators B,  part  to  the  rich  solution  storage  tank  A.  This  is  followed  by  a 
weak  copper  ammonia  solution  as  a  preliminary  wash,  which  follows  the 
rich  solution  down  through  the  charge.  When  this  begins  to  appear  it  is 
turned  into  the  wash-solution  storage  tank  A.  Following  comes  the 
steam  wash.  Steam,  at  a  pressure  of  5  Ib.  per  square  inch,  is  admitted 
above,  and  gradually  works  to  the  bottom,  heating  the  charge  and  con- 
densing to  a  film  of  water  which  displaces  most  of  the  copper-ammonia 
solution.  The  remainder  of  the  ammonia,  volatilized  by  the  steam,  is 
carried  out  with  it  to  be  condensed  in  one  of  the  condensers  D.  When  the 
ammonia  content  of  the  vapors  issuing  from  the  tank  has  fallen  to  0.5 
NHa,  the  wash  is  considered  finished,  the  steam  is  shut  off  and  the  tank  is 
emptied  of  the  exhausted  tailings.  The  steam  washing  takes  about  twenty 


440 


THE  HYDROMETALLURGY  OF  COPPER 


AMMONIA  LEACHING  441 

hours,  and  about  100  Ib.  of  steam  is  used  per  ton  of  charge.  By  this  method 
of  steam  washing  one  can  leach  rich  ores  with  little  loss  of  ammonia  in  the 
tailings  and  without  using  a  large  volume  of  wash-water,  which  would  too 
greatly  dilute  the  distilling  solution. 

There  are  three  units  of  evaporators.  Section  Y,  Y  shows  one  of  these 
units,  the  evaporators  B  B  in  two  stages,  each  provided  below  with  a 
filter  C.  As  the  copper-ammonia-bearing  solution  is  evaporated,  the  vola- 
tile ammonia  vapor  passes  over  to  be  condensed  in  the  condenser  D. 
By  thus  arranging  the  evaporators  in  double  effect  the  ammonia,  con- 
centrated to  at  least  18  per  cent  NHs,  is  obtained.  The  copper  oxide 
falls  out  of  the  solution  as  the  ammonia  leaves  it,  to  the  conical  bottom, 
of  the  evaporator  and  passes  to  the  filters  C,  C.  The  precipitate  con- 
tains as  much  as  80  per  cent  Cu,  the  filtrate  carries  but  0.4  per  cent  Cu  and 
0.9  per  cent  NHs.  This  is  re-treated  in  a  secondary  operation  to  precip- 
itate the  balance  of  the  copper  and  to  produce  a  waste  solution  of  but 
0.025  per  cent  NHs. 

The  ammonia  is  originally  purchased  as  aqua  ammonia  or  ammonia 
hydrate,  but  soon  picks  up  C02,  then  becoming  ammonium  carbonate. 
Due  to  the  large  amount  of  CCb  evolved  in  the  leaching,  vents  are  pro- 
vided in  the  tank  covers. 

It  will  be  noticed  that  if  the  tailings  contain  gold  or  silver  in  such  quan- 
tity as  will  justify  it,  these  metals  may  be  readily  cyanided.  The  tailings, 
acid  free,  are  in  excellent  condition  for  such  treatment. 


CHAPTER  XXXIII 
REFINING  OF  BLISTER-COPPER 

COPPER  REFINING 

Blister-copper,  or  black  copper,  whether  produced  in  the  blast- 
furnace, the  converter,  or  the  reverberatory  furnace,  or  by  melting  the 
concentrate  from  the  native  copper  ore  of  the  Lake  Superior  region,  still 
contains  impurity,  principally  arsenic  with  sulphur  and  iron,  and  all  im- 
purity must  be  removed  by  refining.  If  the  copper  contains  gold  and  silver 
in  quantity  to  warrant  (20  to  40  oz.  silver  per  ton),  it  is  melted  without 
attempting  to  refine,  and  cast  into  anodes  that  then  are  subjected  to  elec- 
trolytic refining.  If  there  is  but  little  precious  metal  hi  the  copper,  it 
may  be  directly  refined  in  the  copper  refining-furnace.  Fig.  241  is 
a  sectional  elevation  and  plan  of  a  40,000  to  60,000-lb.  copper 
refining  furnace.  It  is  14  by  19  ft.  hearth  dimensions,  and  has  a  fire- 
box 5J  by  6^  ft.,  or  30  ft.  area,  and  carries  a  fire-bed  4J  ft.  thick.  The 
hearth,  2J  ft.  deep,  has  a  brick  or  a  sand  bottom.  If  of  sand,  the  bottom 
is  carefully  smelted  in.  Beneath,  the  hearth  is  vaulted  for  ventilation. 
The  bridge,  5  ft  wide,  is  strengthened  by  a  double  conker-plate,  and  on 
either  side  and  in  the  roof  over  the  fire-bridge,  are  ports  that  are  opened 
when  an  oxidizing  flame  is  desired.  In  the  elevation,  at  the  front  end,  is 
to  be  seen  the  outlet-flue  that  leads  to  the  stack  or  chimney.  The  chim- 
ney is  close  to  the  furnace,  but  is  not  shown  in  the  plan.  The  charge  of 
ingots  of  blister-copper  is  put  in  at  the  side  door.  The  door  is  then  tightly 
closed,  and  vigorous  firing  follows.  The  charge  melts  after  several  hours. 
The  front  door  is  then  opened,  and  whatever  slag  has  formed  during  the 
melting  is  skimmed. 

Next  follows  the  rabbling,  the  object  of  which  is  to  oxidize  a  portion  of 
the  copper  and  the  impurities  with  it.  The  operation  years  ago  Consisted 
in  striking  the  surface  of  the  bath  with  a  rabble  in  such  a  way  as  to  splash 
the  metal  and  agitate  it,  thus  exposing  it  to  the  action  of  the  air.  The 
present  way  is  to  insert  a  f-in.  pipe  just  beneath  the  surface  of  the  metal 
and  force  compressed  air  through  it  to  agitate,  and  at  the  same  time  to 
oxidize  it.  The  air-ports  of  the  furnace  also  are  opened  and  the  flame  is 
made  an  oxidizing  one.  The  action  proceeds  to  the  stage  of  "  set  copper," 
CU2O  having  been  by  this  time  formed,  and  in  part  dissolved  in  the  copper. 

4A2 


COPPER-REFINING  FURNACE 


443 


Iron,  sulphur,  and  arsenic  partly  volatilize,  and  partly  oxidize  and  enter 
the  slag  that  is  formed  at  the  same  time ;  this  is  skimmed  off. 

The  copper  oxide  must  be  removed  by  poling.     This  is  a  reducing  action 
in  which  the  air-ports  are  closed  to  give  a  reducing  flame,  and  spruce  or 


poplar  poles  are  inserted  at  the  front  door  into  the  metal.  The  outer  end 
of  the  pole  is  raised  to  force  the  butt-end  beneath  the  surface  of  the  metal. 
At  the  same  time  a  wheelbarrowload  of  charcoal  is  thrown  in  to  cover  the 
surface,  to  exclude  air,  and  to  reduce  cuprous  oxide.  As  the  hydrocarbon 


444  REFINING  OF  BLISTER-COPPER 

of  the  wood  is  evolved  and  the  moisture  evaporates,  that  is,  as  the  wood 
burns,  reduction  takes  place.  The  operation  requires  an  hour  or  two. 
Additional  poles  are  inserted  to  replace  those  consumed.  Samples  of  a 
few  ounces  of  the  copper  are  removed  in  a  small  ladle  from  time  to  time 
and  examined  to  note  the  progress  of  reduction.  The  "  tough  pitch  " 
(the  point  at  which  the  cuprous  oxide  is  completely  reduced  to  metallic 
copper)  is  the  end  in  view. 

The  charge  is  now  ready  for  dipping  or  ladling.  Hand-ladles,  holding 
25-lb.,  or  large  "  bull-ladles  "  holding  200  Ib.  and  carried  by  an  overhead 
trolley  or  crawl,  are  used.  The  dipping  or  molding  consumes  three  hours. 
The  copper  is  kept  hot  by  occasional  firing,  and  by  keeping  the  surface  of 
the  metal  covered  with  charcoal.  The  charcoal  serves  also  to  keep  the 
copper  in  pitch,  or  in  the  condition  of  tough  copper.  The  molds  into  which 
the  copper  is  poured  from  the  ladles  are  of  the  shape  required  by  the  trade. 
There  are  required  ingots  or  bars  suited  to  remelting  for  making  brass; 
wire  bars  of  a  form  convenient  for  rolling  into  wire;  and  rectangular  cakes, 
often  18  in.  square  and  4  in.  thick,  but  also  of  dimensions  giving  2000  to 
4000  Ib.  weight.  The  size  of  the  cakes  is  suited  to  the  size  of  the  sheets  of 
copper  into  which  they  are  to  be  rolled. 

MELTING  AND  REFINING  "  LAKE  "  COPPER 

The  product  from  which  copper  is  made  in  the  Lake  Superior  region  is  a 
concentrate  (called  locally  "  mineral"),  which  averages  70  per  cent 
copper  in  native  form,  accompanied  with  a  self -fluxing  or  fusible  gangue. 
In  addition  there  occur  pieces  of  copper  of  different  sizes,  from  that  -of  the 
fist  to  several  tons  in  weight,  called  mass-copper.  The  small  pieces  are 
handled  easily,  and  are  shipped  to  the  smelting  works  in  barrels.  It  is 
called  barrel-work.  The  larger  pieces  called  "  mass  copper"  or  simply 
"mass,"  are  70  per  cent  copper.  For  the  large  pieces  that  cannot  be 
charged  at  the  side  doors,  a  hatch-opening  with  a  clamped  brick  cover  is 
provided.  Large  pieces  are  raised  by  a  crane  and  charged  through  the 
hatch,  also  concentrate  or  mineral  to  make  a  charge  of  36,000  to  40,000  Ib. 
The  charging  takes  place  immediately  after  the  dipping  and  the  repairing 
or  fettling  the  furnace.  The  furnace  is  now  closed,  and  firing  proceeds  for 
several  hours.  As  the  charge  melts  and  slag  forms,  it  is  skimnted  until 
the  metal  is  completely  melted  and  the  surface  is  clear.  The  operation 
of  refining  then  continues  as  has  been  described  above.  The  slag  contains 
15  to  25  per  cent  copper,  partly  as  entrained  prills  and  flakes,  and  partly 
as  cuprous  oxide. 

This  slag  is  smelted  in  a  blast-furnace  with  added  limestone  to  make 
the  resulting  slag  fusible,  using  anthracite  coal  and  a  portion  of  coke 
for  fuel. 


COPPER-CASTING  MACHINE 


445 


In  recent  practice  in  the  Lake  Superior  country  the  operation  of 
melting  is  performed  in  one  furnace  and  refining  in  another.  A  furnace, 
18  by  40-ft.  hearth  area,  melts  100  tons  of  mineral  of  67  per  cent  copper  hi 
twenty-four  hours,  using  30  tons  of  coal.  The  charge  to  the  second  fur- 


FIG.  242. — Endless  Mold  Casting  Machine. 


FIG.  243. — Endless  Mold  Casting  Machine. 

nace  is  supplied  in  two  portions,  the  second  following  as  soon  as  the  first  is 
melted. 

The  copper  is  poled,  and  when  thus  refined,  is  cast  into  ingots  on  an 
endless-mold  casting-machine,  see  Figs.  242  and  243,  or  on  a  Walker  cast- 
ing-machine, see  Fig.  245.  The  copper  is  tapped  to  casting-machine  over  a 
spout  which  is  hinged,  the  lip  being  raised  or  lowered  to  regulate  the  flow. 


446 


REFINING  OF  BLISTER  COPPER 


In  Fig.  147,  the  endless  chain  receives  the  molten  copper  upon  the  chain 
of  molds  at  a,  and  this  slowly  travels  along  until  the  now  solid  ingot  drops 
upon  a  grating  and  then  on  an  endless  chain  conveyor  in  the  water  box  at  c. 
Here  it  is  thoroughly  cooled,  lifted  to  d,  and  drops  upon  the  floor  at  the 
foot  of  the  slide  e  to  be  removed  for  shipment. 


THE  MAKING  OF  ANODES  AND   OF  COMMERCIAL  CATHODE  COPPER 

Melting  Furnace. — The  melting  down  of  blister-copper  in  ingot  form  is 
performed  in  a  coal-fired  reverberatory  furnace,  40  ft.  long  of  a  capacity 
of  as  much  as  400,000  Ib. 

Charging  Machine. — This,  Fig.  244,  consists  of  a  crane  carrying  a 
transverse  carriage  whereby  an  arm  or  ram  is  introduced  into  the  furnace, 
the  paddle  B  at  the  end  carrying  its  load  of  several  ingots  piled  on 
one  another.  Upon  the  arm  rests  a  racked  bar  that  receives  an  inde- 


FIG.  244. — Clarke  and  Antisell  Charging-crane. 

pendent  movement  by  means  of  the  cylinder  c.  This,  at  the  right 
moment,  shoves  the  load  off  the  paddle  to  drop  upon  the  furnace  bottom. 

The  charge  is  melted  and  poled  to  give  a  smooth  anode,  then  tapped, 
in  regulated  flow. 

The  duty  imposed  upon  these  anode  casting-furnaces  is  extremely 
severe  and  constant.  They  hold  about  200,000  Ib.  of  metal,  and  no 
sooner  has  the  last  anode  of  a  charge  been  cast  than  a  fresh  charge  begins. 
In  order  to  complete  their  duty  of  300,000  Ib.  copper  (1J  charges  per 
twenty-four  hours),  it  is  often  necessary  to  begin  blowing  compressed  air 
into  the  metal  as  soon  as  the  bath  is  sufficiently  deep  to  submerge  the 
air-pipes,  taking  care,  however,  not  to  let  the  oxidation  proceed  too  far 
before  the  entire  charge  is  collected,  else  there  is  danger  of  such  a  violent 
evolution  of  862  gas,  from  the  reaction  between  cuprous  oxide  and  the 
cuprous  sulphide  of  the  fresh  converter  copper,  that  there  is  danger  of  the 
melted  metal  being  blown  out  of  the  furnace.  Two  of  these  furnaces  are 
in  constant  use,  with  a  third  and  larger  one  in  reserve. 


WALKER  CASTING  MACHINE 


447 


The  pigs  contain  about  98.3  per  cent  copper,  and  this  process  brings 
them  up  to  about  99.3  per  cent,  at  which  stage  they  will  cast  into 
smooth  anodes. 


|< 10  's  H  "C.4«f  Bcuiac* »i 


r 


FIG.  245. — Walker  Casting  Machine. 

The  Walker  Casting  Machine. — This,  as  shown  in  Fig.  245,  is  of  the 
horizontal  wheel  type  and  is  capable  of  casting  25  tons  per  hour.  The 
metal  flows  through  the  tapping-slit  at  the  front  end  of  the  furnace  into  a 


448  REFINING  OF  BLISTER  COPPER 

suspended  ladle  B,  from  which  it  is  poured  into  an  anode-mold  at- 
tached to  a  platform  conveyor  operated  hydraulically.  When  the  mold 
is  filled,  the  ladle  is  dropped  to  the  horizontal  position,  and  the  conveyor 
is  moved  so  as  to  bring  the  next  mold  into  position.  The  copper  is  chilled 
by  a  spray  and  when  "  set  "  is  dumped  automatically  from  the  mold  onto  a 
conveyor  operating  through  a  tank  of  water. 

In  the  case  of  cathodes  from  the  tank  house  of  a  refinery,  these  are 
charged  by  the  charging  machine,  Fig.  244,  to  a  large  reverberatory  fur- 
nace, melted  down  and  tapped  to  the  Walker  casting-machine.  The  molds, 
as  shown  in  Fig.  245,  are  making  ingots  for  commercial  use. 


CHAPTER  XXXIV 
ELECTROLYTIC  COPPER  REFINING 

ELECTROLYTIC  COPPER-REFINING  PLANT 

Blister-copper  from  copper  reduction  works  in  the  Western  States, 
as  well  as  that  imported,  is  profitably  treated  by  this  process.  Not"  only 
can  pure  copper  be  produced  from  impure  material,  but  the  gold  and 


FIG.  246. — Ground  Plan  of  Electrolytic  Copper  Refinery. 

silver  in  the  blister  can  be  separated,  parted  and  refined.  The  copper 
from  the  reduction  works,  in  the  form  of  rough  bigots,  or  even  as  anodes, 
is  sampled  and  assayed  to  determine  the  content  of  precious  metal  and  of 
copper. 

Refinery. — The  following  is  the  description  of  a  refinery  of  the  capacity 

449 


450 


ELECTROLYTIC  COPPER  REFINING 


of  50,000  tons  per  annum,  and  covering  22  acres.     It  combines  the  best  of 
the  standard  methods. 

Operation. — Referring  to  Fig.  246,  the  blister-copper  is  received  over 
the  railroad  spur  B,  weighed  in  the  control  weighing  room  E,  sampled  in 
the  sampling  room  F,  and  then  placed  in  storage  under  the  crane-way  L, 
from  which  it  is  taken  as  required,  to  the  furnace  building  J,  K.  This 
has  at  one  end  two  reverberatory  anode  furnaces  for  melting  the  blister, 
and  two  where  the  cathodes  produced  in  the  process  are  remelted  into 


FIG.  247.— Electrolytic  Tank. 

merchantable  form.  The  anode  furnaces  are  lined  with  silica  brisk  and 
are  each  of  100  tons'  daily  capacity.  The  furnaces  are  fired  witn  slack 
coal,  and  have  a  waste-heat  boiler,  so  as  to  utilize  the  steam  for  power  pur- 
poses about  the  plant.  The  raw  material  is  charged  into  the  furnace  on 
one  side  from  the  industrial  railway  by  means  of  a  charging  machine 
(Fig.  244)  and  the  molten  copper  is  tapped  fronVthe  other  side  into  anode 
molds.  The  depth  of  the  bath  of  molten  metal  is  from  21  to  24  in. 
and  during  the  twenty-four  hour  period  the  cycle  of  operations  would 
be:  For  charging,  1J  hours;  for  melting  down,  nine  hours;  for  refining  and 


ELECTROLYTIC  COPPER  REFINING 


451 


poling,  seven  hours;  for  casting,  6i  hours.  The  anodes  are  stored  under 
the  craneway  0,  in  plentiful  supply  to  feed  the  tank-house  U,  where  the 
electrolytic  copper  is  produced. 

The  tank-house  contains  512  electrolytic  tanks  in  two  bays,  a  bay  hav- 
ing 8  sections  of  thirty-two  tanks  each.  Each  tank,  to  hold  28  anodes  and 
29  cathodes,  is  13  ft.  long,  and  of  cross-section  shown  in  Fig.  247. 

The  current  passes  through  each  of  the  30  plates  or  anodes  to  the 
cathodes  placed  between,  and  anodes  and  cathodes  are  2.6  in.  apart. 
Assuming  that  the  anodes  are  2  by  3  ft.  in  size,  we  have,  hi  each  tank,  a 
total  area  of  360  sq.  ft.  through  which  the  current  is  passing  with  a  density 
of  20  amperes  per  square  foot.  We  thus  have  a  total  of  7200  amperes. 
The  pressure  in  passing  through  this  2.6  in.  of  electrolyte  to  the  cathode  is 
0.40  volt.  The  tanks  in  series  would  therefore  give  a  pressure  of  102  volts. 


+ 
-JWV 


Buss  bar  loss  divided  into  ft 
st  of  12  or  more  taaka 


-Contact  Ion 


Contact  at  cathode  connecti 


A 


FIG.  248. — Current-flow  Walker  Multiple  System. 

Fig.  248  represents  the  arrangement  of  anodes  and  cathodes  in  the 
Walker  multiple  system.  It  will  be  seen  that  the  flow  of  current  through 
each  vat  is  hi  parallel  but  from  vat  to  vat  in  series. 

The  building  must  be  kept  at  a  uniform  temperature  of  about  80?  F., 
causing  a  tendency  in  cold  weather  to  "  sweating  "  on  roof  and  walls,  on 
account  of  the  large  amount  of  moisture  evaporated  from  the  electrolytic 
tanks.  For  this  reason  the  heating  should  preferably  be  by  the  circulation 
of  dry  hot  air  to  absorb  the  evaporated  moisture. 

The  copper  electrolyte  consists  of  about  4  per  cent  copper  hi  the  form  of 
sulphate,  together  with  about  12  per  cent  of  free  acid.  During  the  deposi- 
tion of  the  copper  the  electrolyte  becomes,  in  course  of  time,  polluted  by 
the  impurities  contained  in  the  anode  copper,  on  account  of  which  a  certain 
amount  of  the  electrolyte  is  periodically  drawn  off  and  treated  in  the  regen- 
erating plant,  the  purified  solution  being  returned  to  the  copper  electrolytic 
tanks.  The  rate  of  deposition  of  copper  on  the  cathode  will  be  governed 


452  ELECTROLYTIC  COPPER  REFINING 

by  the  current  density,  which  should  not  amount  to  more  than  20  amperes 
per  square  foot  of  anode  surface  for  copper  carrying  up  to  100  oz.  of  gold 
and  silver  to  the  ton.  Should  the  amount  of  precious  metals  be  appreci- 
ably less,  the  current  density  may  be  increased  up  to,  say,  30  amperes, 
thereby  increasing  the  speed  of  deposition  and  shortening  the  time  of 
operation.  Should  the  deposition  take  place  too  quickly,  there  is  danger 
of  occluding  precious  metals  with  the  cathode  copper.  A  cycle  of  opera- 
tions in  the  tanks  at  the  20-ampere  current  density  would  occupy  from 
three  to  four  weeks. 

Starting  Sheets. — One  section  of  the  tank  house  is  devoted  to  the  prepa- 
ration of  cathode  starting  sheets,  for  which  purpose  a  copper  plate  coated 
on  each  side  with  oil  or  graphite  is  used  as  a  cathode  blank,  and  a  thin  layer 
of  copper  deposited  thereon  by  electrolysis.  This  is  stripped  off  from  each 
side  and  forms  the  cathode  starting  sheet.  In  the  routine  operation  the 
electrolytic  tanks  are  first  charged  with  anodes,  and  then  hung  with 
cathode  starting  sheets,  and  the  current  is  started.  It  is  kept  on  for  a 
prescribed  number  of  days,  varying  from  ten  to  twelve ;  then  the  cathodes 
which  have  been  formed  on  the  starting  sheets  are  removed,  the  solution 
is  lowered  in  the  tanks,  the  anodes  are  taken  out  temporarily,  and  the 
slime  is  removed  and  sent  to  the  slime  refinery.  The  anodes  are  then 
replaced,  fresh  cathode  starting  sheets  hung,  and  the  current  is  turned  on 
for  another  period  of  ten  to  twelve  days;  then  the  "  pulling  "  is  repeated, 
except  that  upon  this  occasion  what  is  left  of  the  anodes  is  removed  and 
goes  back  as  scrap  to  be  remelted  in  the  anode  furnaces.  It  will  be  noted, 
therefore,  that  for  each  anode  going  to  the  tanks  two  cathodes  of  lighter 
weight  are  formed.  The  theoretical  deposition  of  the  copper  would  be 
approximately  0.062  Ib.  of  copper  per  ampere  day,  and  a  current  efficiency 
of  at  least  90  per  cent  should  be  acquired.  The  lecessary  pressure  would 
be  about  0.4  volt  per  tank,  and  upon  this  basis  the  rate  of  deposit  should 
amount  to  about  6  Ib.  of  cathode  copper  per  kilowatt-hour. 

The  cathodes  are  weighed  on  leaving  the  department,  and  stored,  from 
which  point  they  are  fed  to  the  refined  copper  furnace  for  melting  into  the 
shape  required  by  the  consumer,  for  example,  ingots,  wire  bars,  and  cakes. 

The  Slime  Refinery  TF,  Fig.  246. — The  precious-metal  slime  reclaimed 
from  the  electrolytic  tanks  amounts  to  about  30  tons  per  month,  opntain- 
ing  about  440,000  oz.  of  gold  and  silver.  The  slime  is  first  freed  of  the 
electrolyte  by  settling  until  it  contains  about  50  per  cent  moisture.  It  is 
then  boiled  with  sulphuric  acid  to  remove  soluble  copper,  after  which  it  is 
put  through  a  filter  press  and  roasted.  This  roasted  product  is  combined 
with  a  portion  of  unroasted  slime  to  make  up  a  furnace  charge.  The  result 
will  be  the  melting  of  about  20  tons  per  month  of  slime  containing  from 
15  to  25  per  cent  moisture,  the  melt  taking  place  in  a  small  reverberatory 
furnace,  the  product  of  which  is  dore*  bars. 


HANDLING  OF  ANODES  AND  CATHODES 


453 


Parting.— Dore*  bars  form  the  anodes  for  an  electrolytic  deposition,  which 
takes  place  in  an  electrolyte  slightly  acidified  with  nitric  acid,  the  silver 
being  deposited  on  cathode  blanks  in  crystalline  form,  and  the  gold  set- 
tling as  a  black  mud  containing  about  equal  quantities  of  gold  and  silver. 
The  resulting  silver  crystals  are  scraped  off  the  cathode  blanks,  washed  and 
melted  into  fine  silver  bars.  The  gold  mud  is  purified  by  treatment  with 
nitric  acid  to  part  the  gold  from  the  silver,  the  solution  going  back  to  the 
silver  electrolytic  tanks,  and  the  go'd  sand  melted  with  borax  into  fine 
solid  bars. 


FIG.  249. — Mechanical  Handling  of  Anodes  and  Cathodes. 

Mechanical  Handling  of  Anodes  and  Cathodes. — In  Fig.  249,  the  view 
above  shows  the  anodes,  and  below,  the  cathodes,  the  entire  anode  or 
cathode  contents  of  a  tank  being  lifted  and  transferred  in  one  operation. 
The  frame  shown  is  brought  down  by  a  traveling  crane  over  the  tank,  all 
the  anodes  or  all  the  cathodes  of  the  tank  hooked  on  to  it,  the  frame  and 
its  load  lifted  and  transferred  to  a  clear  space  at  the  end  of  the  tank 
house  where  the  fragments  of  anodes  can  be  removed  to  be  sent  to  the 
anode  furnace,  while  the  built-up  cathodes  are  stored,  to  go  later  to  the 
refined-copper  furnace. 

The  Regenerating  Plant  7,  Fig.  246. — The  best  proportion  of  acid  and 


454  ELECTROLYTIC  COPPER  REFINING 


copper  for  the  electrolyte  is  10  per  cent  EbSC^,  and  15  per  cent 
(equivalent  to  6  per  cent  Cu).  When  the  copper  exceeds  this  quantity, 
the  resistance  increases;  hence  copper  is  removed  from  the  circulating 
electrolyte  or  solution  if  in  excess,  to  bring  the  amount  to  the  required 
proportion,  the  quantity  of  iron,  arsenic,  antimony,  and  tellurium  gradually 
increases,  and  a  time  comes  when  the  electrolyte  becomes  foul  with  them 
and  the  excess  must  be  removed.  Antimony  can  be  kept  low  by  the  daily 
addition  of  a  small  amount  of  salt,  which  precipitates  as  an  oxychloride. 

To  purify  the  electrolyte  the  following  method  is  used.  A  portion 
of  the  electrolyte  is  diverted  in  a  constant  flow  to  tanks  reserved  for  the 
purpose  of  purification.  These  have  insoluble  sheet-lead  anodes  and 
copper  cathodes.  A  strong  current  is  used,  so  that  not  only  is  copper 
deposited,  but  also  the  impurity.  The  deposit  collects  loosely  upon  the 
copper  plates  and  falls  to  the  bottom  of  the  tanks.  Every  two  months 
the  accumulated  mud,  containing  40  to  60  per  cent  copper,  is  cleaned  out 
and  reduced  in  a  reverberatory-refining  furnace  to  form  impure  bars  of 
copper.  The  purified  electrolyte  is  returned  to  the  main  system. 

Circulation  of  the  Electrolyte.  —  To  avoid  short-circuiting,  and  to 
increase  the  activity  and  regularity  of  deposition,  the  electrolyte  is  made 
to  flow  or  circulate  through  the  tanks,  entering  the  top  of  each  tank  near 
the  end,  passing  downward  between  the  plates,  and  finally  rising  and 
flowing  away  through  an  overflow  pipe  at  the  other  end.  After  the  solu- 
tion has  flowed  through  two  tanks  in  this  way,  it  enters  a  launder  that 
returns  it  to  the  collecting  or  sump-tank.  Thus  every  pair  of  tanks  has 
an  independent  circulation.  The  sump-tank  receives  all  the  electrolyte, 
to  be  here  heated  by  means  of  a  steam-coil  to  40°  C.,  the  effect  of  the 
warming  being  to  decrease  the  electrolytic  resistance.  It  is  then  pumped 
up  to  a  distributing  or  stock-tank,  and  once  more  enters  the  circulation. 

Testing  the  Current.  —  Besides  the  voltmeter  and  ammeter  to  be  found 
at  the  switch-board  in  the  power-house,  it  is  customary  to  use  a  voltmeter 
for  constantly  testing  the  drop  in  potential  between  the  anodes  and 
cathodes.  For  this  a  forked  rod  is  used,  which  touches  the  two  plates 
and  takes  a  small  current  through  a  portable  voltmeter.  A  slight  drop  of 
pressure  indicates  short-circuiting. 

The  Power-house,  Q,  Fig.  246.  —  Here  are  installed  five  motor  generator 
sets  each  to  give  a  constant  current  of  5000  amperes  at  60  to  115  volts. 
Four  of  these  are  arranged  to  work  in  groups  of  two  in  parallel  directly 
to  the  two  separate  circuits  in  the  tank-house.  The  fifth  unit  is  connected 
to  act  as  a  spare  for  any  of  the  others. 

Cathode  Storage.  —  Leaving  the  tank  house  as  heavy  cathodes,  these 
are  placed  under  the  craneway  0,  thence  delivered  to  the  reverberatory 
refining  furnaces  J,  there  to  be  melted  and  cast  in  the  forms  required  by 
the  consumer. 


REFINING  AND  OPERATING  COSTS  455 


CAPITAL  REQUIREMENTS 

Not  only  is  capital  invested  in  the  buildings  and  the  equipment  of  the 
plant,  but  it  is  required  for: 

(1)  The  stock  of  anodes  in  process  of  treatment. 

(2)  The  stock  of  anodes  awaiting  treatment. 

(3)  The  copper  constantly  contained  in  electrolyte. 

(4)  The  copper  needed  for  the  heavy  conductors  transmitting  the 
current. 

The  result  of  this  large  demand  upon  capital  is  to  restrict  the  operation 
of  plants  to  places  near  financial  centers,  like  New  York,  where  cheap  money 
is  available,  the  copper  near  the  market,  and  labor  abundant.  These  con- 
siderations may  outweigh  the  advantages  of  having  the  plant  near  cheap 
water-power. 

COST  OF  REFINERY  AND  OPERATING  COSTS 

Cost  of  Refinery. — To  refine  50,000  tons  of  blister-copper  annually 
containing  gold  and  silver,  the  output,  consisting  of  refined  copper  in  the 
shape  of  wire  bars,  cakes,  and  ingot,  of  gold  as  fine  gold  bars,  and  silver 
as  fine  silver  bars,  the  construction  costs  will  be  as  follows: 


CONSTRUCTION  COSTS 

Furnace  building $  67,500 

Two  anode  units 95,000 

Two  refined-copper  units 92,200 

Four  cranes,  two  services,  two  charging 40,600 

Auxiliary  equipment 84,800 


Total  furnace  department $380,600 

Tank  house  building $127,000 

Sixteen  32-tank  sections  and  circulation  system  and  equip- 
ment   258,000 

Electrolytic  circuit  conductors,  etc 96,000 

Four  service  cranes 25,000 

Auxiliary  apparatus 51,000 


Total  tank  house  department 557,000 

Power-house  building  and  crane. .  . : $  42,000 

Five  electrolytic  motor  generator  sets  and  equipment 116,000 

Motor-driven  pumps,  etc 32,200 

Auxiliary  apparatus 30,200 


Total  power-house  department 220,400 


456 


ELECTROLYTIC  COPPER  REFINING 


Slime  refinery  building $  38,000 

Settling  and  boiling  tanks,  filter  press  and  roaster 16,500 

Furnace  and  flue  system 22, 100 

Electrolytic  parting  cells  and  equipment 18,800 

Gold-refining  equipment 2,600 

Auxiliary  apparatus 11,700 

Total  slime-refining  department $109,700 

Shops  building  and  equipment $  31,000 

Office  and  laboratory  and  warehouse  equipment 44,000 

Sampling  apparatus 7,500 

Control  weighing  scales  and  housing 7,400 

Storage  cranes,  craneways,  and  locomotive  crane 36,700 

Receiving  and  shipping  apparatus 3,600 

Industrial  locomotives  and  cars 47,800 

Industrial  tracks  and  railroad  sidings 24,500 

Miscellaneous  auxiliary  apparatus 7,600 

Bluestone  and  acid  necessary  to  make  up  electrolyte 25,000 

Total  miscellaneous $235,100 

Total $1,502,800 

These  figures  are  the  more  valuable  that  they  show  in  what  unusual 
ways  the  money  has  to  be  spent,  causing  costs  to  mount  rapidly. 

Operating  Costs. — These  are  as  below  estimated,  viz. : 


Labor. 


Material. 


Total 
Per  Ton. 


Anode  making ,. . 

Electrolytic  refining. . 

Refined-copper  casting 

Slime  refining 

General  expenses 

Interest  on  investment  (5  per  cent) 


$1 . 155 

2.025 

1.00 

.463 

.926 


$1.213 

2.464 

1.304 

.461 

.464 

1.537 


$2.368 
4.489 
2.404 
.924 
1.390 
1.537 


Totals. 


$5.669 


$7.443* 


$13.112 


*  Per  ton  of  refined  copper. 

SCHEDULE  OF  COPPER  AND  COPPER  ORE  PRICES 

In  Utah,  custom  copper  smelting  works  pays  $19  an  ounce  for  the  gold; 
for  95  per  cent  of  the  silver,  based  on  the  New  York  quotation,  and  for 
copper,  after  deducting  1  per  cent  from  the  wet  assay,  a  further  deduction 
of  3  to  4  per  cent  from  the  New  York  price  per  ounce.  A  charge  of  $5  per 
ton  is  made  for  treatment  and  when  the  insoluble  exceeds  40  per  cent  then 
5  cents  per  unit  for  all  over  this.  Zinc  in  excess  of  10  per  cent  is  charged 
for  at  the  rate  of  30  cents  per  unit.  The  deduction  of  3  or  4  cents  from 
the  copper  price  is  to  cover  the  cost  of  freight  and  refining. 


PRICES  OF  COPPER  AND  COPPER  ORES  457 

Copper. — The  New  York  price  is  expressed  in  cents  per  pound,  quota- 
tions being  given  for  Lake  copper  cast  in  the  form  of  cakes  for  rolling  into 
sheets,  into  ingots  for  remelting  to  make  castings,  brass,  and  bronze,  or  for 
wire-bars  for  drawing  into  wire.  Casting-copper  is  not  as  pure  as  that  which 
is  to  be  rolled  into  sheets  or  drawn  into  wire.  Electrolytic  copper  is  made 
by  remelting  cathodes  (the  product  of  electrolytic  refining)  into  ingots, 
cakes,  or  wire-bars.  Cathodes  are  held  at  £  cent  less  than  electrolytic 
copper,  the  difference  paying  for  the  remelting. 

In  London  copper  is  sold  by  the  long  ton  in  English  money,  and  is  of 
various  brands.  Standard  copper,  formerly  called  g.m.b.  (good  merchant- 
able bars),  is  the  grade  upon  which  the  others  depend.  Besides  these 
brands  we  have: 

"  English  tough  copper,"  "  best  selected  "  or  "  standard."  If  sold 
for  immediate  delivery  it  is  called  "spot  copper";  if  the  customer  will 
take  it  at  the  expiration  of  three  months  it  is  called  "  three-months  copper." 

It  is  the  business  of  dealers,  and  others  interested  in  copper,  to  keep 
statistics  of  the  supply  of  available  copper,  which  is  called  the  "  visible 
supply."  When  this  is  small  the  price  naturally  rises,  and  the  reverse  is 
true  when  it  is  large. 


PART  VI 
LEAD 


CHAPTER   XXXV 
PROPERTIES  OF  LEAD  AND  ITS  ORES 

Physical  Properties  of  Lead. — This  metal,  so  soft  that  its  purity  may 
be  judged  of  by  scratching  with  the  thumb  nail,  is  commercially  divided 
into  corroding,  common,  and  antimonial  lead.  Of  the  base  meals  it  is  the 
heaviest,  having  a  specific  gravity  of  11.37.  It  soon  tarnishes  on  a  freshly 
cut  surface,  becoming  dull  gray.  It  is  malleable,  so  that  it  can  be  rolled 
into  sheets  and  pressed  into  pipe,  but  it  has  little  tenacity.  It  freezes  at 
325°  C.  and  boils  at  1525°  C. 


CHARACTERISTICS  OF  LEAD  ORES 

Classes  of  Lead  Ores. — The  lead  ores  are  those  in  which  lead  is  the 
principal  constituent.     The  term  is  applied  also  to  mineral  aggregates 
consisting  of  more  than  10  per  cent  lead.     The  lead  ores  may  be  divided  f 
into  two  classes,  thfe  sulphide  and  the  oxidized.     The  terms  are  used  only  < 
according  to  the  constituent  that  is  in  excess;  in  many  lead  ores  both  sul- 
phides and  oxides  are  found.     Ore  containing  no  lead  is  called  dry,  and 
when  carrying  lead,  leady.     The  latter  term  is  the  opposite  of  dry,  but  we 
do  not  term  a  leady  ore  a  wet  one. 

Galena. — Pure  galena  contains  86.6  per  cent  lead  and  13.4  per  cent  • 
sulphur.     In  nature  it  occurs  with  gangue  or  vein-matter.     When  there  is 
much  of  the  latter  it  can  readily  be  concentrated.     The  following  table 
gives  an  idea  of  the  lead-content  of  ore,  before  and  after  dressing: 

GALENA   ORES 


Locality. 

RAW  ORE. 

CONCENTRATE 

Pb, 
Per  Cent. 

Pb, 
Per  Cent. 

Ag.  oi.. 
Per  Ton. 

S.  E.  Missouri  . 

4.3 

80.0 
62.0 

12.0 
80.0 
0.3 

30.0 
19.8 
30.0 

Minnie  Moore,  Wood  River,  Idaho 

Rockville,  Wis 

St.  Joseph,  Mo  
Kellogg,  Idaho 

7.0 
11.0 
10.0 

70.0 
60.0 
55.0 
49.5 

Col  Sellers,  Leadville,  Colo 

Coeur  d'Alene 

461 


462 


PROPERTIES  OF  LEAD  AND  ITS  ORES 


Galena  from  the  Mississippi  Valley  contains  little  silver.  Galena  concen- 
trate from  S.  E.  Missouri  contains  Pb,  69  per  cent;  SiO2,  1.4  per  cent; 
Fe,  5.1  per  cent;  CaO,  3  per  cent;  Zn,  0.8  per  cent;  S,  15.5  per  cent. 
That  from  the  Rocky  Mountain  region  is  not  only  argentiferous,  but  may 
contain  gold.  The  precious  metals  as  well  as  the  lead  determine  the  value. 
Metallic  sulphides,  such  as  pyrite  and  blende,  are  often  associated  with 
galena,  and  with  the  gangue  may  carry  so  much  of  the^old  and  silver  that 
concentrating  leads  to  a  serious  loss  of  the  metals  and  is  omitted.  If  by 
hand-picking  ore  can  be  brought  to  contain  30  to  40  peFcent  lead,  it  is  a 
desirable  ore  for  the  smelter.  When  of  this  tenor  in  lead,  and  free  from 
other  sulphides,  it  carries  but  5j)er  cent  sulphur  and  needs  no  preliminary 
roasting,  and  is  smelted  directly. 

Oxidized  Lead  Ores. — Little  lead  oxide  is  found  in  nature.  The  ores 
classed  here  under  oxidized  ores  are  the  result  of  the  alteration  of  galena. 
They  include  the  carbonate  (cerussite)  and  the  sulphate  (anglesite)  of  lead. 
The  minerals  are  mixed  with  metallic  oxides  and  vein-matter  or  gangue  in 
nature,  and  when  sandy  or  earthy,  the  ore  is  called  sand  or  soft  carbonate, 
and  when  hard  and  stony,  hard  carbonate.  In  many  deposits  we  find  ore 
that  originally  was  galena,  profoundly  altered  to  cerussite  or  anglesite. 
The  subjoined  table  gives  the  composition  of  some  of  the  so-called  car- 
bonates : 


CARBONATE    ORES 


Locality. 

Pb, 
Per  Cent. 

Si02, 
Per  Cent. 

Fe, 
Per  Cent. 

CaO, 
Per  Cent. 

S, 
Per  Cent. 

Ag.  oz. 
Per  Ton. 

Southwest  Missouri  

72.0 

Leadville,  Colo  

29.0 

11.6 

24.3 

5.0 

Leadville,  Colo  

21.0 

22.5 

18.2 

2.4 

0.9 

65.0 

Red  Mountain,  Colo  

18.4 

41.6 

11.4 

1.7 

1.8 

128.0 

Eureka,  Nev  

33.2 

3.0 

24.1 

1.1 

2.0 

27  5 

Bingham,  Utah  

51.5 

12.5 

2.6 

3.2 

6.0 

21.1 

Horn  Silver  mine,  Frisco,  Utah 

50.0 

15.2 

3.4 

0.5 

8.3 

78.3 

Of  the  ores  of  the  table,  that  from  Eureka,  Nev.,  contains  4.2  per  cent 
of  arsenic,  which  forms  an  arsenical  speiss  when  smelted.  Tlje  Horn 
Silver  ore,  apparently  oxidized,  has  the  lead  in  the  form  on  inglesite 
(PbS04),  and  matte  is  formed  from  it  in  smelting.  In  oxidized  ores  the 
silver  is  apt  to  occur  as  a  chloride;  the  gold  probably  is  native. 

There  are  many  lead  minerals,  but  those  not  mentioned  occur  in  small 
quantity  and  are  not  considered  among  the  commercial  lead  ores.  • 


SMELTING  ON  THE  ORE-HEARTH  463 


THE  SMELTING  OF  LEAD-BEARING  ORES 

When  lead  concentrates  are  to  be  smelted  only  for  the  lead  content, 
as  is  done  in  parts  of  the  Mississippi  Valley,  a  simple  plant  with  a  rever- 
beratory  furnace,*  or  the  American  ore-hearth,  is  sufficient.  In  the 
rocky  Mountain  region  the  lead  ore  is  not  smelted  to  recover  only  the  lead. 
The  lead  of  the  ore  is  employed  as  a  collector  of  the  gold  and  silver  of  other 
ores  that  are  smelted  at  the  same  time.  By  use  of  the  ore  hearth,  a  large 
part  of  the  lead  is  recovered  cheaply  and  simply,  but  a  part  is  lost  in  the 
resultant  slag.  In  silver-lead  smelting  it  is  essential  that  the  slag  be 
comparatively  free  from  lead  and  consequently  from  silver. 

SMELTING  ON  THE  ORE-HEARTH 

The  ore-hearth  cannot,  as  regards  capacity,  or  cost  per  ton  of  ore 

treated,  be  used  hi   silver-lead   smelting.     It  can  be  quickly  started  or 

^  stopped,  and  put  in  operation  with  little  cost  for  fuel,  so  it  well  serves  the 

purpose  of  extracting  the  lead  from  small  amounts  of  low-silver  ore,  from 

time  to  time,  by  the  men  who  themselves  have  mined  the  ore. 

The  Hearth. — Fig.  250  represents  a  sectional  elevation,  a  front  eleva- 
tion of  the  lower  part,  and  a  plan  of  an  American  ore-hearth.  It  consists 
of  a  cast-iron  pan  or  crucible  a,  2  by  2J  ft.  by  1  ft.  deep,  to  contain  a 
bath  of  lead.  The  back  p  and  the  two  sides  n,  n,  above  the  crucible,  are 
water-cooled  castings.  The  blast  from  a  fan-blower  (not  shown)  enters  by 
the  tuyere-pipe  6  through  the  back  at  o.  At  g  is  a  sloping  cast-iron  plate 
called  the  work-stone,  and  at  t,  a  pot,  placed  to  recover  the  lead  that  flows 
down  over  the  work-stone.  The  pot  is  kept  hot  by  a  wood  fire  below. 
The  structure  is  surmounted  by  a  brick  top  to  receive  and  carry  off  the 
fumes. 

Operation. — By  means  of  the  blast,  a  glowing  coal  fire  is  made  that 
fills  the  crucible  of  the  hearth,  a;  residue  from  the  previous  run,  con- 
taining metallic  lead,  and  15  to  20  Ib.  galena,  not  finer  than  pea-size,  is 
spread  over  the  fire.  The  charge  soon  becomes  red  hot,  and  the  lead,  set 
free,  finds  its  way  to  the  crucible  at  the  bottom.  More  ore  is  then  added, 
and  the  material  in  the  hearth  is  pried  up  gently  with  a  bar  to  keep  the 
mass  open  and  hot  throughout.  Lumps  form,  and  are  drawn  out  on  the 
work-stone,  g,  and  gray  slag  that  forms  at  the  same  time  is  separated 
and  the  rich  residue  returned  to  the  hearth.  Ore  and  fuel  are  again  added, 

*  The  treatment  of  lead  ore  in  reverberatory  furnaces  has  not  made  headway  in  the 
United  States.  There  are  two  reasons  for  this:  In  the  silver-lead  districts,  the  ore  has 
not  been  of  sufficient  grade  in  lead  to  warrant  the  treatment,  and  lead  ore  has  been  in 
great  demand  as  a  collector  to  mix  with  other  ores.  Secondly,  in  the  Mississippi  Valley 
where  silver-free  lead  concentrate  is  made,  the  question  of  skilled  labor  for  reverberatory 
.furnace  work  has  had  an  influence. 


464 


PROPERTIES  OF  LEAD  AND  ITS  ORES 


15  to  20  Ib.  at  a  time,  and  operations  continue  until  lead  fills  the  crucible, 
while  on  the  top  floats  the  fuel,  unreduced  ore,  and  half-fused  material. 
One  man  with  a  bar  at  intervals  loosens  and  stirs  the  charge,  raising  it 
slowly,  while  another  with  a  shovel  draws  upon  the  work-stone  the  half- 
fused  mass  floating  on  the  lead.  Here  he  separates  and  rejects  the  gray 
slag,  and  returns  the  rich  residue  to  the  charge.  A  fresh  charge  is  then 
added,  and  the  work  progresses  in  the  manner  described.  The  lead  over- 


VERTICAL  SECTION  ON  LINE  C-D 


FRONT  ELEVATION 


HORIZONTAL  SECTION  ON  LINE  A-R 


FIG.  250. — American  Ore-hearth. 

flows  the  crucible  and  runs  down  a  groove  made  in  the  work-stone  into  the 
kettle,  i.     When  the  kettle  fills,  the  lead  is  skimmed  and  ladled  into.molds. 
With  air  in  excess  the  red-hot  galena  is  in  part  oxidized  to  PbSO-i, 
in  part  to  PbO.     Both  these  react  upon  the  galena  thus: 

(1)  PbS+2PbO  =3Pb+SO2. 

(2) 

Also  the  glowing  fuel,  reacting  upon  any  remaining  oxide,  completes  the 
reduction. 


THE  NEWNAM  HEARTH 


465 


To  operate  the  ore-hearth,  a  blower  and  power  to  run  it  are  needed. 
Much  lead  is  volatilized,  and  so  the  treatment  is  not  suited  to  argentif- 
erous galena.  The  gray  slag  that  is  produced  still  contains  35  to  40  per 
cent  lead,  and  is  sold  to  smelting-works.  The  direct  recovery  of  the 
lead  is  75  to  85  per  cent,  the  higher  figure  having  been  obtained  in  recent 
practice. 

The  Newnam  Hearth. — A  recent  development  of  the  ore-hearth  has 
been  the  introduction  of  mechanical  rabbling,  thus  doing  away  with  the 


^^ 


FIG.  252.  —  Newnam  Hearth  (end  view  of  plant). 

most  laborious  and  hottest  work  of  the  hearth,  that  is,  the  rabbling  of  the 
charge.     The  increase  in  its  size  and  capacity  is  another  advantage. 

Fig.  251  shows  such  a  rabbling  machine,  which  is  electrically  trav- 
ersed upon  an  overhead  track,  free  from  possible  obstruction.  The 
machine  carries  a  rabble  or  poker,  which  plows  its  way  -through  the  crust 
of  the  molten  bath  from  the  left  to  the  right  end  of  the  furnace,  a  distance 


466  PROPERTIES  OF  LEAD  AND  ITS  ORES 

of  8  feet.  Following  it  is  a  furnace  helper,  who,  with  a  long-handled 
shovel,  removes  the  gray  slag  and  pushes  back  unfused  lumps,  while  the 
furnaceman  adds  a  thin  layer  of  ore  and  a  little  coal,  as  in  the  American 
ore-hearth  as  just  described.  The  rabble,  seen  in  251,  sloping  downward 
into  the  bath,  is  now  lifted,  and  the  machine  returned  to  the  left  end  of  the 
furnace  to  make  another  stroke. 

Fig.  252  is  a  cross-section  of  the  plant  with  the  many  details  plainly 
marked,  the  goose-neck  rising  above  the  building  carries  the  dust  and  fume 
to  a  balloon  flue,  where  the  material  is  withdrawn  into  a  car. 


CHAPTER  XXXVI 
SILVER-LEAD  SMELTING 

SILVER-LEAD   BLAST-FURNACE   SMELTING 

This  is  a  blast-furnace  method  of  treatment,  applicable  to  a  great 
variety  of  ores  containing  lead,  silver,  gold,  and  even  copper.  By  it,  ores 
containing  the  precious  metals  with  no  lead,  are  treated  with  lead-bearing 
ores,  thus  using  the  lead  of  one  ore  as  the  collector  of  the  gold  andosilver  of 
another.  This  is  the  most  effective  method  of  treating  such  ores.  The 
precious  metals  are  extracted  from  the  ore  by  a  blast-furnace  treatment, 
using  lead-bearing  ores,  carbonaceous  fuel,  and  flux. 

Oxidized  ore  can  be  directly  smelted  in  the  blast-furnace,  bub  sulphide 
ore  is  first  roasted.  Methods  of  roasting  are  described  in  the  chapter  on 
Roasting. 

The  ore  to  be  smelted  is  charged  into  the  blast-furnace  as  in  iron  or 
copper  smelting,  with  a  calculated  quantity  of  flux,  which,  for  lead  ore,  is 
iron  ore  and  limestone.  The  precaution  is  taken  to  use  lead-bearing  ore 
enough  to  make  the  lead  content  of  the  charge  at  least  10  per  cent.  It 
has  been  found  that  if  a  smaller  proportion  of  lead  than  this  is  used,  the 
precious  metals  are  not  so  well  collected  in  the  base-bullion,  or  work-lead, 
produced  in  the  smelting.  To  a  charge,  as  above  constituted,  is  added 
10  per  cent  or  more  of  coke,  not  on]y_to  melt  the  charge,  but  to  reduce  the 
lead  oxide  to  metal  and  the  ferric  oxide  to  the  ferrous  form. 

RECEIVING,  SAMPLING    AND  BEDDING  OF  LEAD  ORES 

Where  ore  is  treated  in  a  small  way  for  the  recovery  of  the  lead,  as  in 
Missouri,  no  particular  provision  is  made  for  storage.     In  various  custom 
smelting  works  in  the  Rocky  Mountain  region,  where  lead  ores  are  treated 
with  others  by  methods  of  silver-lead  smelting,  and  where  ores  are  bought 
outright  for  treatment,   the  handling  becomes   complicated.    A   plant 
treating  ore  from  its  own  mine  is  called  a  mine's  works,  and  here  less  ' 
attention  is  given  the  sampling  and  storing  ^bTfhe  ore.     In  a  custom  works,  , 
therefore,  ores  of  many  kinds  are  received,  some  containuig^lead,  sonic 
having  little  lead  but  carrying  silver  and  gold. 

467 


468 


SILVER-LEAD  SMELTING 


BEDDING  ORES  AT  A  CUSTOM  WORKS 

The  ore  is  received  in  lots  of  a  few  tons  up  to  those  of  several  car- 
loads. Each  lot  is  separately  weighed,  sampled  as  described  in  the  chap- 
ter on  Sampling,  then  assayed,  and  purchased.  If  different  kinds  of  ore 
were  smelted  separately  the  process  would  involve  endless  change  and 
labor,  and  so  it  has  become  the  custom  to  "  bed  "  the  ore  in  large  bins 
or  stalls,  each  holding  several  hundred  tons.  When  so  bedded  the  mix- 
ture, called  a  "  mix  "  is  treated  as  a  single  ore  The  different  kinds  of 


FIG.  254.— Ore-bed. 


ore  are  unloaded  separately  into  the  bin,  and  each  kind  is  spread  out  in  an 
even  layer  before  the  succeeding  one  is  added,  as  indicated  in  Fig.  254. 
When  the  ore  is  to  be  used,  shoveling  is  done  at  the  floor  and  all  parts 
above  fall  down  and  mix,  since  a  steep  face  of  ore  is  constantly  maintained. 
Thus  a  uniform  mixture  of  the  different  ores  is  obtained  for  smelting.  The 
contents  of  the  bed  are  treated  in  the  books  of  the  company  as  a  single 
ore. 

Side  Ores. — Besides  ore  bedded  in  this  way,  large  lots  may  be  kept 
separate.  Even  smaller  lots,  of  which  a  moderate  amount  is  to  be  used  per 
charge,  may  be  so  kept.  Such  are  called  "  side  ores." 

Charging  from  Feed-bins. — Ore  and  fluxes  are  often  stored  in  hopper- 
bottom  bins  or  pockets,  to  be  drawn  off  in  weighed  quantity  into  the  fur- 


BEDDING  ORES 


469 


nace   charge-car.     Fig.  253   shows  such   an  installation.     The  charge  is 
dropped  into  the  furnace  by  opening  the  double-hinged  bottom  of  the  car. 

The  coke  and  fluxes  are  stored  separately  in  large  piles,  so  that  in  case 
of  failure  of  railroad  delivery 
due  to  washouts,  etc.,  the  fur- 
nace shall  not  have  to  close 
down. 

Charging  Fuel. — Coke  for 
the  charge  is  forked  into  the 
coke-buggies,  using  a  fork 
with  lj-in.  spaces,  thus  leav- 
ing fines  which  are  generally 
thrown  away.  Such  loss  may 
amount  to  5  per  cent. 

A  bed,  formed  of  10  to  15 
per  cent  SiC>2,  20  to  28  per  cent 
Fe,  and  20  to  28  per  cent  Pb, 
roasts  well.  Mixtures  contain- 
ing less  lead  and  more  pyrite 
than  this  roast  readily,  but  a 
mixture  pulverulent,  when 
roasted,  tends  to  make  more 
flue-dust  in  the  blast-furnace, 

while  with  the  proportion  of  lead  above  specified  it  sinters  and  makes  a 
desirable  lumpy  product. 

Sintering  Ores. — Besides  sintering  by  blast-roasting,  the  Dwight- 
Lloyd  machine,  described  on  page  112,  has  come  extensively  into  use,  not 
only  for  silver-lead  ores,  but  for  fine  iron  ores  as  well. 


FIG.  253. — Charging  Feed-car. 


GENERAL  ARRANGEMENT  OF  A   SMELTING  WORKS 

A  One-furnace  Smelting  Plant. — In  Fig.  256  we  illustrate  a  com- 
pletely assembled  works.  Since  the  materials  pass  downward  from  level  to 
level  this  is  called  a  side-hill  or  terraced  plant.  To  suit  present  prac- 
tice the  reverberatory  roasting  furnaces  would  be  replaced  by  a  multi- 
pie  hearth-roaster  and  a  D wight-Lloyd  sintering  machine  occupying  but 
one-third  the  space. 

The  30-in.  Dwight-Lloyd  machine  has  treated  60  tons  of  sulphide 
ore,  or  80  tons  where  non-sulphide  ores  are  mixed  in.  The  42-in.  machines 
will  yield  80  to  120  tons  under  similar  circumstances,  and  at  higher  speeds 
200  tons  and  over. 

Fig.  255  is  a  cross-section  of  a  Dwight-Lloyd  sinter  plant,  con- 
taining 42-in.  machines.  An  inclined  conveying  belt  delivers  charge  to  a 


470 


SILVER-LEAD  SMELTING 


horizontal  one  on  the  top  floor  which  has  a  tripper  to  deliver  it  to  the  feed 
hoppers.  At  the  right  it  discharges  to  a  railroad  car  receiving  the  sintered 
ore.  From  the  building  a  9-ft.  balloon  flue  leads  away  the  gases  to  a  stack. 
An  exhaust  fan,  one  to  each  furnace,  draws  these  away  to  the  stack. 

Composition  of  the  Charge  for  Sintering. — The  sinter  charge  may  be 
quite  variable.  It  is  customary  to  rough-roast  ore  high  in  sulphur  until 
it  is  reduced  to  about  12  to  15  per  cent  in  sulphur.  A  satisfactory  charge 
would  contain  13  per  cent  Pb,  23  per  cent  Si02,  25  per  cent  Fe  plus  Mn,  12 
per  cent  sulphur.  This  rough-roasted  product  is  blast-roasted  down  to 
3  per  cent  sulphur. 

Ore  in  cars,  which  has  to  be  sampled  or  crushed  for  roasting,  enters  by 
the  track  a'  on  the  extreme  left,  or  otherwise  by  the  track,  j,  Fig.  256.  At  the 
latter  track,  while  unloading,  every  tenth  shovelful  can  be  retained  in  the 


SECTION  B-B  ^ n/_ 

FIG.  255. — Cross-section  of  Sinter  Building. 

arc  while  the  nine-tenths  is  bedded  upon  the  mixing  floor.  The  car  can  then 
be  sent  to  the  sampling  mill  by  track  a'.  Crushed  ore  for  the  roasters  is 
received  in  a  car  &,  then  trammed  to  the  pile  d.  Here  it  is  withdrawn,  as 
needed,  and  taken  to  the  hoppers  e',  e'}  of  the  two  reverberatory  roasters 
in  the  roaster-building  H.  A  flue  -X",  common  to  them  both,  leads  to  the 
stack  /.  The  roasted  ore  is  taken  by  wheelbarrows  to  the  cooling  flpor  R, 
then  to  any  bin  V  of  the  mixing  floor.  Non-roasting,  sampled  ^ore  is 
trammed  by  the  tracks  s',  t,  for  bedding  at  any  bin  V.  Fuel  and  fluxes 
are  also  unloaded  to  the  mixing  floor.  Charges,  as  fast  as  assembled  and 
weighed,  are  raised  by  the  "  lift  elevator  "  to  the  blast-furnace  feed-floor. 
The  blast-furnace  is  supplied  with  air  by  a  rotary  blower  w,  driven  by  a 
steam  engine  in  the  "  engine-blower  room." 

.  The  gases  from  the  blast-furnace  pass  away  to  the  flue  p  by  the  goose- 
neck down-take  o,  thence  to  the  dust-chamber  /  and  to  the  stack  g.     The 


SILVER-LEAD  SMELTING  WORKS 


471 


base-bullion  is  loaded  into  the  car  n,  at  the  front  of  the  furnace  building. 
The  slag,  flowing  through  the  fore-hearth  m,  is  taken  by  slag-pots  to  the 
edge  of  the  dump  and  there  emptied. 


THE  SILVER-LEAD  BLAST-FURNACE 

This  differs  from  the  copper  blast-furnace  principally  in  the  crucible 
which  in  the  copper  blast-furnace  is  above  the  floor  level  on  a  carriage 
while  in  the  silver-lead  furnace  the  crucible  rests  upon  the  floor,  having  a 
deep  cavity,  the  crucible  proper. 


472 


SILVER-LEAD  SMELTING 


Fig.  257  is  a  perspective  view  and  Fig.  258  shows  elevations  of  a 
furnace,  giving  many  details  of  construction.  The  brick  crucible,  heavily 
bound  with  plates  and  steel  rails,  has,  as  shown  in  the  transverse  section,  a 
channel  called  a  "  lead  well,"  widening  upward  from  the  bottom  of  the 
crucible  and  having  a  spout  where  the  molten  lead  (which  fills  crucible 
and  lead-well  alike)  may  be  removed.  There  are  six  water-jackets  on 


FIG.  257. — Perspective  View  of  Silver-lead  Blast  Furnace. 


each  side  and  one  at  each  end  (see  Fig.  258),  forming  the  furnace 
bosh.  Another  tier  of  jackets  forms  the  lower  part  of  the  shaft,  while  its 
upper  part  of  brick  reaches  to  the  charge  door.  On  each  long  side  are 
two  counter-weighted  doors.  When  a  furnace  is  to  be  fed  the  doors  are 
opened,  one  at  a  time,  to  do  so.  Each  lower  side  jacket  is  pierced  with  two 
tuyere  openings,  making  twenty-four  tuyeres  in  all,  there  being  none  in  the 


SILVER-LEAD  BLAST-FURNACE 


473 


474 


SILVER-LEAD  SMELTING 


end  jackets.  The  jackets  are  held  in  place  by  a  frame  of  I-beams  that 
bear  against  them.  There  is  a  main  water-supply  pipe  whence  proceeds 
a  feed-pipe  to  each  jacket.  The  spill  from  the  jackets  is  taken  to  a  waste 
trough,  set  above  the  bustle-pipe.  The  bustle-pipe  surrounds  the  furnace 
on  three  sides. 

The  tuyeres  themselves  are  well  shown  in  Fig.  259.  Each  one  has 
its  own  gate-valve  so  that  it  can  be  shut  off  for  any  desired  purpose. 

Air  is  supplied  by  the  bustle-pipe  e,  Fig.  261.  Iron-pipe  connections 
are  now  preferred,  as  in  Fig.  259. 

In  the  case  of  the  closed  top  the  gases  pass  away  to  the  dust-flue  p  of 
Fig.  256  by  way  of  the  down-take  o,  which  slopes  at  an  angle  of  45°; 


Asbestos  Gasket 

in  machined 
Retaining  Groove 


Rubber-packed 
Tuyere-box  Cover 
(Anaconda  Type) 


Detachable  Tuyere-box: 
Fastening  in  Open  Lug 

fusible  or  Wooden  Plug 


lylo.r  Patent,  Welded  Tuyere 

WORKING   POSITION 


1 


SWUNG  BACK  TO  REMOVE  JACKET 


FIG.  2,.9.  —  Blast-furnace  Tuyere. 

so  that  the  flue-dust  will  not  lodge  in  the  pipe.  At  Fig.  261  the  down- 
take  is  flatter,  but  it  can  be  cleared  by  a  rabble  worked  from  the  feed- 
floor. 

The  arrangement  of  the  water-cooled  tap-  jacket,  as  shown  in  Fig.  262, 
is  an  end  jacket,,  having  a  bosh.  Beneath  this  is  set  the  cast-iron  water- 
cooled  tap-jacket  t,  12  in.  square.  This  is  wedged  beneath  and  on  either 
side  with  brick  and  fireclay,  and  the  breast  is  bricked  up.  The  conical 
aperture  or  tap-hole  is  plugged  with  clay.  When  sufficient  slag  has  accu- 
mulated within  the  furnace,  say  every  fifteen  minutes,  a  hole  is  pierced 
through  the  clay  of  the  tap-hole  and  the  slag  flows  out  over  the  spout  s, 
being  received  in  a  slag-pot  or  into  a  fore-hearth. 


OPEN-TOP  BLAST-FURNACE 


475 


OPEN-  AND  CLOSED-TOP  BLAST-FURNACES 

In  the  blast-furnace  building,  Fig.  256,  and  in  Fig.  257,  are  views  of 
a  closed-top  furnace,  while  in  Fig.  261  we  have  sections  of  an  open-top 
one.  In  the  former  the  smoke  is  caught  in  the  closed-top  and  passes  away 
overhead  and  the  furnace  can  be  fed  to  the  level  of  the  feed  floor;  in  the 
second  the  gases  pass  away  by  the  down-take  x  which  limits  the  charge 
level  to  its  bottom.  On  the  other  hand  the  whole  furnace  is  accessible 


180 


FIG.  261. — Lead  Blast  Furnace. 

from  above  both  for  dumping  charges  and  for  cleaning  out  the  furnace  after 
shutting  down. 

The  Open-top  Furnace.— The  furnace,  Fig.  256,  is  fed  by  hand, 
shoveling  the  charge  through  the  feed-doors  into  the  furnace.  At  large 
smelting  plants  this  method  has  been  superseded  by  "  automatic  charging  " 
(see  Fig.  261).  To  prepare  for  this  the  closed  top  above  the  feed- 
floor  is  omitted.  Then  for  charging  a  large  car  having  a  drop-bottom  is 
brought  to  the  furnace  and  the  contents  of  the  car  dropped  into  it.  Such 
a  furnace  is  shown  at  Fig.  261.  On  the  left  we  have  a  half  section,  half 
elevation,  transverse  to  the  furnace;  at  the  right  there  is  shown  on  one 
half,  a  longitudinal  section,  and  beside  it  a  fore-hearth.  On  the  right  half 


476  SILVER-LEAD  SMELTING 

is  the  longitudinal  elevation  and  the  brick  down-take  x,  also  a  cross  sec- 
tion of  the  down-take. 

The  furnace  is  45  by  160  in.  at  the  tuyere  level,  and  widens  to  95  by 
160  in.  at  the  top.  There  is  a  single  steel  jacket  at  each  end,  and  two 
boshed  jackets  at  each  side.  It  will  be  noticed  that  the  shaft  widens  out 
as  it  ascends.  There  are  twenty  tuyeres,  ten  on  each  side,  the  openings 
in  the  jackets  being  3J  in.  diameter.  Each  tuyere  has  its  own  metal  blow- 
pipe and  shut-off  valve.  The  down-take  x  is  3  ft.  3  in.  by  7  ft.  11  in. 
inside  dimensions,  and  leads  the  smoke  to  a  flue  common  to  several  fur- 
naces. When  a  furnace  is  to  be  charged  a  light  sheet-iron  cover  at  the 
floor  level  is  readily  removed  'by  hand.  At  3  ft.  below  the  charge-floor 
are  set  transversely  angle-iron  "  spreaders,"  so  that  the  materials,  falling 
from  the  car,  shall  be  spread  or  distributed  evenly.  In  this  particular 
installation  the  charge-car,  of  about  the  same  width  and  length  as  the 
furnace,  approaches  it  at  the  side;  in  other  cases  at  the  end  of  the  furnace. 

OPERATING  THE  BLAST-FURNACE 

Blowing  In. — The  crucible  and  lead-well  are  dried  out  and  warmed  by  a 
wood  fire  for  twenty-four  hours.  They  are  then  cleaned  out  and  the 
crucible  is  filled,  using  the  well-cleaned  lead  ingots  from  the  drossing 
kettles.  These  are  piled  in  cross  order,  filling  the  crucible.  On  them  is 
laid  a  wood  fire  as  high  as  the  top  of  the  breast,  and  after  closing  this  the 
fire  is  lit  at  the  tuyeres.  The  wood  ablaze,  more  dry  wood  is  added,  then 
coke  to  the  depth  of  2  ft.  with  a  little  iron  ore  and  limestone,  just  enough 
to  flux  the  coke-ash.  Then  begins  the  charging,  at  first  using  double  the 
usual  quantity  of  coke,  then,with  the  furnace  full,  dropping  to  the  usual 
proportion. 

The  blast  is  increased  gradually  during  one  to  two  hours,  until  the  fur- 
nace is  in  full  operation.  As  the  slag  accumulates  it  is  tapped,  while  the 
lead,  accumulating  in  the  crucible  and  lead-well,  is  either  dipped  from 
the  lead-well  with  ladles,  or  tapped  through  a  lead  tap-hole  at  the  level 
of  the  top  of  the  crucible.  The  amount  of  lead  removed  at  one  time  is 
limited  to  1000  to  1500  Ib.  to  keep  the  crucible  always  full.  The  opera- 
tion of  filling  and  starting  takes  seven  hours. 

Regular  Work  on  the  Charge-floor. — In  a  small  furnace  this  consists 
in  wheeling  ore  and  fuel  from  the  bins  to  the  charge-scales,  weighing  the 
required  amounts  into  charge-barrows  or  charge-cars,  then  dumping  the 
charge  thus  prepared  upon  the  feed-plates  in  front  of  the  furnace  doors. 
Every  material  except  the  foul  slag  is  weighed,  and  even  the  slag  is  added 
by  shoveling  in  regular  amounts.  A  charge  is  dumped  on  one  charge-plate 
and  the  coke  for  it  on  the  other.  The  coke  is  fed  in  an  even  layer,  and  the 
plate  thus  cleared  receives  an  ore-charge.  The  ore-charge  on  the  other 


FRONT  OF  BLAST-FURNACE 


477 


charge-plate  is  then  added,  taking  care  to  place  the  larger  ore  at  the  middle 
of  the  furnace  and  the  fine  at  the  walls  and  corners.  Such  a  distribution 
should  be  made  as  to  cause  the  smoke  and  gas  to  rise  evenly  throughout. 
The  blast  tends  to  ascend  at  the  walls  more  than  in  the  middle,  but  by  the 
distribution  it  is  compelled  to  rise  evenly.  The  plate,  now  cleared  of  the 
ore-charge,  receives  the  coke  that  is  next  to  go  in.  The  content  of  the  fur- 
nace remains  at  the  level  of  the  charge-floor,  new  charges  being  supplied 
as  the  surface  sinks. 

Regular  Work  at  the  Ground  or  Slag-floor. — This  consists  in  regulating 


FIG.  262. — Front  of  Silver-lead  Blast  Furnace. 


the  water-supply  at  the  jackets,  seeing  that  the  tuyeres  are  clean  and  open, 
tapping  and  stopping  the  slag,  and  when  the  slag  and  matte  are  separated 
in  a  fore-hearth,  tapping  the  matte,  placing  the  slag-pots  (see  Fig.  262), 
and  removing  them  to  the  dump  when  full. 

Front  End  of  Furnace. — We  show  at  Fig.  262  the  front  end  of  a  large 
blast-furnace.  Immediately  at  the  front  is  to  be  seen  the  fore-hearth,  d, 
6  ft.  by  9  ft.  in  size,  mounted  on  wheels,  and  here  the  separation  of  the 
matte  from  slag  is  effected.  The  top  of  the  slag  in  the  fore-hearth  soon 
becomes  crusted  over  with  a  slag  crust,  but  beneath  this  is  the  molten 


478 


SILVER-LEAD  SMELTING 


slag  and  matte.  The  heavier  matte  settles  out  while  the  lighter  slag  escapes 
by  a  spout  at  the  front  of  the  fore-hearth  into  the  bowl  of  a  large  slag-car, 
g,  mounted  on  wheels.  There  is  a  tap-hole  on  the  side  opposite  that  shown, 
where  the  fore-hearth  is  tapped  in  order  to  remove  the  matte  which  then 
accumulates.  This  matte  is  generally  received  into  flat  or  dish-shaped 
bowls  and,  when  solidified,  it  is  removed  by  a  traveling  crane.  It  is  broken 
by  hammers  and  is  then  ready  for  crushing  to  3-mesh  size  for  roasting. 

When  a  slag-bowl  is  filled  it  is  removed  by  an  electric  motor  to  the  damp. 
A  shell  or  "  skull "  of  solidified  slag  has  formed,  lining  the  interior  of  the 
bowl  and  the  top.  A  hole  is  broken  in  the  top  and  the  molten  contents 
poured  out  as  shown  in  Fig.  263.  This  shell  is  returned  to  the  feed- 
floor  to  add  to  the  charge,  since  it  contains  enough  lead  and  silver  to  make 
this  worth  while.  Of  late  some  smelting  works  reject  it  all,  saying  that  it 
contains  so  much  zinc  as  an  impurity  that  it  does  not  pay  to  return  it. 


FIG.  263. — Side-dumping  Slag-pot. 

Slag  is  also  conveniently  removed  by  means  of  large  slag  pots  mounted 
on  trucks  and  drawn  by  an  industrial  locomotive,  see  Fig.  155,  where  the 
firct  two  trucks  are  carrying  shallow  matte-bowls,  the  farther  two  slag 
bowls. 

Dressing  Base-bullion. — When  handling  base-bullion  from  the  small 
blast-furnace  of  50  to  100  tons  capacity  of  a  generation  ago,  the  ordinary 
custom  was  to  tap  it  when  full  from  the  lead  well,  into  a  cooler,  a  pot  of  one 
ton  capacity.  The  dross  of  the  metal  would  form  a  crust  on  top,1  and  by 
aid  of  a  perforated  skimmer  was  put  into  molds,  the  clean  lead  being  then 
added  on  top  to  yield  a  smooth-looking  bar.  The  refinery  got  the  copper- 
containing  dross,  and  so  did  not  complain. 

The  present  method,  using  large  furnaces,  is  to  dross  the  lead  as  follows: 
The  lead  is  tapped  into  pots  on  wheels,  and  this  is  accumulated  in  the  dross- 
ing-kettle  until  it  is  filled  with  the  molten  lead.  It  is  now  skimmed,  using 
a  Howard  press,  the  skimming,  after  pressing,  being  dumped  upon  a  chute, 


CHEMICAL  REACTION  OF  THE  BLAST-FURNACES  479 

where  it  is  broken  up  and  returned  to  the  blast-furnace.  The  hot  molten 
lead  is  allowed  to  cool  to  the  casting  temperature  and  siphoned  into  molds, 
as  described  on  page  .  The  resultant  lead,  well  freed  from  copper,  still 
retains  all  its  antimony,  say  1.5  per  cent,  and  this  is  sent  to  the  refinery. 
The  product  is  quite  even,  and  its  assay  is  accurate. 


CHEMICAL    REACTIONS    AND    PHYSICAL    CHANGES    OF   THE 
BLAST-FURNACE 

The  surface  of  the  charge  should  look  dead,  showing  no  visible  heat  or 
flame  (over-fire).  With  much  over-fire,  there  is  a  loss  of  lead  due  to 
volatilization.  The  moisture  in  the  charge  soon  dries  at  the  temperature 
of  the  rising  gases  (200°  C.).  The  heat  thus  absorbed  is  small  and  by  cal- 
culation it  is  found  to  be  but  one-thirtieth  of  the  total  supplied  by  the  fuel 
when  5  per  cent  moisture  is  in  the  charge.  As  the  materials  of  the  charge 
descend  in  the  furnace,  carbon  dioxide  begins  to  be  driven  from  the  lime- 
stone, and  the  iron  reduces  from  the  ferric  to  the  ferrous  form.  Half 
way  down  at  a  temperature  of  800°  C.  the  reactions  are  complete.  The 
lead  in  oxidized  form  is  reduced  by  the  CO  of  the  gas  and  by  the  red-hot 
coke.  Galena  reacts  with  the  iron  oxide  and  carbon  as  follows: 

(3)  PbS+FeO+C  =  Pb+FeS+CO. 

Sometimes  scrap  iron  is  added  to  the  charge,  and  acts  with  galena  or  other 
sulphides  as  in  the  nail  assay  for  silver,  as  follows : 

(4)  PbS+Fe  =  FeS+Pb. 

As  the  lead,  thus  reduced,  drops  through  the  charge,  it  collects  the  gold 
and  silver  as  well  as  a  part  of  the  arsenic  and  antimony,  and  enters  the 
crucible  as  base-bullion.  When  antimony  is  present  it  is  reduced  like  lead 
and  alloys  with  the  latter.  Anglesite  is  reduced  by  contact  with  the  fuel 
and  iron  oxide  according  to  the  following  reaction : 

(5)  PbSO4+FeO+5C  =  Pb+FeS+5CO. 

Where  much  anglesite  is  in  the  charge,  more  than  the  usual  quantity  of 
fuel  is  demanded.  Since  the  affinity  of  sulphur  for  copper  is  greater  than 
for  iron,  copper  sulphide  remains  unreduced,  and  copper  as  an  oxide  takes 
sulphur  from  the  charge  and  with  the  iron  forms  matte.  Lead  to  the 
extent  of  10  to  20  per  cent,  either  as  sulphide  or  in  metallic  form,  is  also 
taken  up  by  the  matte.  To  some  extent  zinc  sulphide  also  enters  the  matte. 


480  SILVER-LEAD  SMELTING 

The  ferrous  oxide,  not  needed  to  satisfy  the  matte,  and  the  CaO  and  MgO 
in  the  charge  must  be  present  in  sufficient  quantity  to  form  a  suitable 
fusible  slag  with  the  silica.  Oxidized  arsenic  compounds  react  with  iron 
and  carbon  producing  a  speiss,  often  of  the  form  Fe4As,  and  require  extra 
fuel.  The  reaction  is  as  follows: 

(6)  As2O3+8FeO+  11C  =  2Fe4As+  11CO. 

The  molten  products  separate  at  the  hearth  according  to  the  specific 
gravity,  that  of  lead  being  11.5,  speiss  6.0,  matte  5.2,  and  slag  3.6.  The 
lead,  collected  in  the  crucible,  is  withdrawn  at  the  lead-well.  The  slag, 
matte,  and  speiss  are  drawn  off  at  the  slag  tap-hole  on  the  level  of  the  top 
of  the  crucible  and  of  the  lead.  The  separation  of  matte  from  slag  is  gen- 
erally effected  in  a  fore-hearth  outside  the  furnace. 

SLAGS   IN   SILVER-LEAD   SMELTING 

The  object  of  silver-lead  smelting  is  to  reduce  the  lead,  and  incidentally 
the  gold  and  silver,  from  the  ore.  The  sulphur  present  forms,  with  the 
copper,  iron  and  a  part  of  the  lead,  a  complex  artificial  sulphide  (matte), 
while  basic  flux  is  added  to  form  a  slag  of  the  composition  that  experience 
shows  necessary. 

Slags  are  silicates  of  extraordinary  complexity;  and  not  all  merely 
fusible  slags  work  well  in  a  silver-lead  blast  furnace.  Type  slags  are  those 
so  proportioned  in  silica,  iron  oxide,  and  lime  as  to  work  well  in  a  blast 
furnace.  Slags  that  vary  from  the  proportions  become  defective  in  opera- 
tion. To  fulfill  the  requirements  of  good  slag,  it  should  have,  in  the  normal 
operation  of  the  furnace,  not  more  than  0.7  per  cent  lead,  or  0.5  oz.  silver 
per  ton,  when  producing  base-bullion  not  higher  than  300  oz.  silver  per  ton. 
The  density  should  not  be  greater  than  3.6.  It  should  not  permit  accre- 
tions to  form  at  the  hearth,  nor  the  creeping-up  or  appearance  of  over-fire. 
If  a  slag  varies  from  one  of  the  types  given,  it  is  either  poorly  reduced  or 
makes  other  trouble  in  the  furnace.  Thus  a  slag  of  the  three-quarter  type, 
in  which  the  CaO  falls  to  20  per  cent  (the  other  constituents  being  given  in 
the  table),  is  found  to  contain,  for  example,  1  per  cent  lead,  and  more  than 
1  oz.  silver  per  ton;  that  is,  it  is  "  dirty."  It  easily  may  happen  that  a 
dirty  slag  is  fusible,  but  we  know  that  the  slag  will  work  satisfactorily  if 
correct  according  to  the  type,  the  other  conditions  of  good  running  being  in 
evidence.  Indeed,  it  is  a  common  experience  that  a  furnace,  working 
poorly  on  an  incorrect  slag,  begins  to  run  well  when  a  correct  slag  comes 
down.  Following  we  give  a  table  of  type-slags  that  have  been  found  to 
work  well  in  practice : 


SLAG  IN  SILVER-LEAD  SMELTING 


481 


TABLE  OF  TYPICAL  SLAGS 


Type. 

SiOJt 
Per  Cent. 

Fe(Mn)O, 
Per  Cent. 

Ca(Ba,Mg)O, 
Per  Cent. 

Quarter-slag 

c 

28 

50 

12 

Silicious  Quarter-slag 

H 

32 

47 

11 

Half-slag 

E 

30 

40 

20 

Half-slag                              

J 

31 

38 

21 

Silicious  half  -slag  
Three-quarter-slag  
Silicious  three-quarter-slag 

I 
F 
M 

35 
33 
36 

38 
33 
31 

17 
23 
23 

Whole  or  1  to  1  slag 

G 

35 

27 

28 

According  to  the  ratio  of  CaO  to  FeO  the  slag  is  called  a  "  quarter,"  a 
"  half,"  or  a  "  one-to-one  "  slag,  etc.  Thus  the  slag  E  of  the  subjoined 
table  is  called  a  half-slag,  the  CaO  being  but  half  of  the  FeO.  The  slag 
C  is  a  quarter-slag,  the  CaO  being  quarter  of  the  FeO.  In  this  table  of 
typical  well  tested  slags  the  three  elements  SiO2,  Fe(Mn)O,  and 
Ca(Mg,Ba)O  are  calculated  to  comprise  90  per  cent.  If  the  sum  varies 
from  this,  the  ratio  is  still  to  be  preserved. 

Since  any  of  the  slags  of  the  table  can  be  used,  the  question  arises, 
which  is  to  be  chosen?  In  this  we  are  guided  by  the  economic  conditions. 
If  the  ore  of  the  district  is  silicious,  and  most  profit  is  derived  by  treating 
the  ore  at  hand,  use  a  silicious  slag  that  needs  the  smaller  amount  of  flux. 
If  irony  or  limy  ores  are  plentiful  and  profitable  to  smelt,  we  use  them, 
substituting  them  for  flux.  It  is  found,  however,  that  slags  of  the  type 
M  and  G  of  the  table  drive  more  slowly  and  require  more  fuel  than  the  basic 
ones.  Perhaps  the  most  satisfactory  of  the  slags,  and  the  one  that  can 
be  used  where  silicious  ores  are  plentiful,  is  the  three-quarter  slag,  F. 
When  a  slag  of  a  certain  type,  for  example  a  silicious  one,  is  not  working 
well  and  is  forming  accretions  in  the  furnace,  a  radical  change  to  a  basic 
type  is  found  beneficial,  or  from  a  basic  slag  to  a  silicious  one. 

ACTION  OF  VARIOUS  BASES  IN  SLAGS 

Iron. — Iron  ore  is  quickly  reduced  to  ferrous  form  under  the  action 
of  the  CO  in  the  furnace  or  of  the  highly  heated  fuel  thus: 


(7) 


Fe2O3+CO  =  2FeO+CO2. 


Iron  oxide,  being  a  stronger  base  than  lead  oxide,  replaces  it  in  the  slag, 
and  the  latter  is  reduced  by  carbon  to  metallic  lead. 


(8) 


PbSiOa+FeO+C  =  FeSiO+Pb+CO.- 


482  SILVER-LEAD  SMELTING 

Manganese. — The  equivalent  for  manganese  is  55,  and  for  iron  56, 
and  they  are  reckoned  as  having  equal  values  for  fluxing.  Manganese  is 
found  in  some  of  the  Leadville  iron  ores  to  the  extent  of  10  to  15  per  cent, 
and  since  by  introducing  another  element,  it  adds  to  the  complexity  of  the 
slag,  it  also  adds  to  the  fusibility. 

The  Alkaline  Earths. — Lime,  magnesia,  and  baryta  act  in  inverse  ratio 
to  their  atomic  weights  in  fluxing  silica,  hence  to  obtain  the  equivalent  in 
lime,  the  percentage  of  magnesia  is  multiplied  by  1.4  and  of  baryta  by 
0.4.  A  slag,  high  in  lime  and  consequently  low  in  iron  like  the  last  three 
in  the  table,  is  of  low  specific  gravity.  Its  use  thus  results  in  a  better  sepa-  & 
ration  of  slag  from  the  heavier  matte.  Lime  being  a  stronger  base  by  one- ' 
half  than  iron,  and  generally  a  cheaper  flux,  the  tendency  is  to  choose  the 
limy  slags.  It  is  noticed  that  the  higher  the  silica  content  of  the  slags  of 
the  table  the  higher  is  the  lime,  and  that  high  silica  calls  for  high  lime. 
Dolomite,  having  a  high  content  in  magnesia,  generally  is  avoided  in 
silver-lead  smelting,  for  it  tends  to  make  slag  pasty  and  streaky,  and  the 
unfavorable  effect  is  aggravated  when  zinc  is  also  present.  Two  analyses 
of  limestone  and  of  dolomite  are  given  below  to  show  conditions  typical 
of  actual  practice. 

Canyon  City  Limestone. — CaO,49.8  per  cent;  MgO,  3.0  per  cent;  SiO, 
3.1  per  cent;  Fe,  0.8  per  cent. 

Iron  County,  Missouri,  Dolomite. — CaO,26.6  per  cent;  MgO,  17.6  per 
cent;  SiO2,  5.1  per  cent;  Fe,  3.3  per  cent. 

Fluorspar. — This  has  no  unfavorable,  but  rather  a  favorable  effect 
upon  the  quality  of  the  slag.  The  fluorine,  however,  uses  CaO,  and  hence 
the  slag  must  analyze  higher  in  CaO  than  the  type  requires,  or  it  will  not 
be  clean. 

Alumina. — It  is  uncertain  whether  alumina  acts  as  an  acid  or  a  base. 
It  is  sufficient  for  the  purpose  of  silver-lead  smelting  to  regard  it  as  a 
neutral  constituent  that  dissolves  in  slag  and  acts  in  neither  way. 

Zinc. — Either  blende  or  zinc  oxide  causes  difficulties  in  the  blast-furnace, 
the  blende  being  the  more  objectionable.  Blende  is  in  part  decomposed 
in  the  presence  of  iron  to  zinc  oxide,  but  the  zinc  in  any  form  tends  to  make 
a  stiff,  pasty,  difficultly  fusible  slag.  It  may  be  regarded,  like  alumina,  as 
being  dissolved  in  the  slag.  It  goes  into  both  the  slag  and  the  ma^te  and 
diminishing  the  specific  gravity  of  the  latter  it  causes  a  less  perfect  separa- 
tion of  the  two.  Where  much  zinc  is  in  the  charge,  it  is  customary 
to  modify  the  type-slag  by  calculating  the  zinc  oxide  as  replacing  one- 
half  the  percentage  of  lime.  Take,  for  example,  the  half  slag  J  of  the 
table. 

In  the  first  column  we  write  the  slag  as  the  type  requires.  In  the  second 
column  we  add  the  8  per  cent  Zn  and  reduce  the  lime  by  4  per  cent  by  which 
the  total  becomes  94.  Since  the  constituents  should  amount  to  but  90 


NATURE  OF  SLAG  AS  AFFECTED  BY  BASES 


483 


Without 
Zinc, 
Per  Cent. 

With 
Zinc, 
Per  Cent. 

Recalcu- 
lated 
Zinc. 
Per  Cent. 

SiOj                                        

31 

31 

29  5 

FeO                                              

38 

38 

36  0 

CaO  

21 

17 

16.0 

ZnO                                              

8 

7  5 

90 

94 

90.0 

per  cent,  all  are  reduced  proportionately  in  the  third  column  so  as  to  give 
90  per  cent  as  the  sum. 

Copper. — Copper  present  in  the  charge  enters  the  matte  when,  as 
generally  is  the  case,  sulphur  is  present  with  which  it  can  combine.  In 
smelting  carbonate  or  oxidized  ores,  which  furnish  no  sulphur,  the  copper 
becomes  reduced,  and  enters  the  base-bullion,  giving  a  lead  so  drossy 
sometimes  as  to  clog  the  lead-well,  and  accumulate  and  solidify  in  the 
crucible.  The  remedy  is  to  supply  sulphide  to  form  matte  into  which  the 
copper  can  enter. 

Antimony. — Either  as  an  oxide  or  a  sulphide,  antimony  is  reduced  like 
lead.  It  alloys  with  the  base-bullion,  making  it  hard,  and  is  removed  and 
recovered  later  in  refining  the  base-bullion. 

Arsenic. — This  frequently  is  encountered  in  silver-lead  smelting. 
When  present  in  small  quantity  it  is  volatilized,  but  in  large  quantity  it 
forms  a  speiss.  Where  it  is  intended  to  produce  a  speiss,  iron  is  provided 
with  which  the  arsenic  unites.  In  the  fire-assay  of  arsenic-bearing  lead 
ores,  a  bead  of  speiss  is  found  attached  to  the  lead  button.  From  the 
percentage  of  this  we  can  compute  the  weight  of  the  speiss  that  will  be 
formed;  and  we  may  assume  that  70  per  cent  of  it  is  Fe.  Where  a  direct 
determination  of  arsenic  is  made  we  can  compute  the  weight,  and  multiply 
this  by  2.3  to  express  the  quantity  of  Fe  to  be  provided  on  the  charge  for 
the  purpose. 

FUEL  IN  SILVER-LEAD  SMELTING 

The  fuels  used  in  silver-lead  smelting  are  coke,  charcoal,  or  a  mixture  of 
the  two.  Wood  and  hard  coal  have  been  used  experimentally,  the  former 
in  certain  cases  of  scarcity  of  fuel. 

Coke. — Coke  is  the  kind  of  fuel  commonly  used.  The  ash  varies 
from  10  to  22  per  cent,  and  the  fixed  carbon  from  89  to  77  per  cent.  In 
coke  of  high  ash,  not  only  is  the  ash  to  be  smelted,  but  the  carbon  is  cor- 
respondingly low,  so  that  the  coke  is  less  efficient.  A  great  difficulty 
with  high-ash  coke  is  that  it  is  often  friable,  making  accretions  or  scaffolds. 
Analyses  of  two  typical  samples  of  coke  give  the  following  results: 


484  SILVER-LEAD  SMELTING 

Connellsville  coke  contains  fixed  carbon  87.5  per  cent,  ash  11.3  per  cent 
and  sulphur  0.7  per  cent;  El  Moro  coke,  fixed  carbon  77.0  per  cent,  ash 
22.0  per  cent  and  sulphur  (when  the  coke  is  made  from  unwashed  coal) 
0.9  per  cent. 

In  computing  a  charge  the  coke-ash  is  taken  into  account,  analyses 
being  as  follows:  Ash  of  Connellsville  coke:  SiO2,  44.6  per  cent;  Fe,  15.9 
per  cent;  CaO,  7.0  per  cent;  MgO,  1.9  per  cent;  ash  of  El  Moro  coke: 
SiC>2,  84.5  per  cent  and  Fe,  5.0  per  cent. 

Charcoal. — This  fuel  is  used  in  districts  far  from  railroads,  where  the 
cost  of  coke  is  high.  It  is  a  good  fuel  for  oxidized  ores,  but  is  friable  and 
makes  undesirable  fine  which  may  form  accretions  or  scaffolds  in  the  fur- 
nace. It  renders  a  charge  more  open  than  coke,  and  contains  less  than 
2  per  cent  ash.  Coke  weighs  25  Ib.  and  charcoal  10  Ib.  per  cubic  foot 
when  loose,  the  weight  of  a  bushel  of  charcoal  being  14  to  16  Ib.  Even 
where  charcoal  is  cheap  it  is  desirable  in  operating  the  furnace  to  use  part 
coke  which,  fed  to  the  walls,  burns  more  slowly  than  charcoal  and  makes 
the  tuyere-zone  hotter  and  gives  a  more  liquid  slag. 

Quantity  of  Fuel.— This  varies  according  to  the  nature  of  the  charge, 
and  generally  is  from  10  to  15  per  cent.  Charges  that  contain  sulphur  and 
make  matte  need  less  fuel  than  oxidized  ores.  Only  sufficient  is  used  to 
give  adequate  reduction  and  a  hot  slag;  and  the  metallurgist  is  guided  by 
these  requirements  in  adding  the  fuel. 

Pulverized  Coal. — This  is  now  being  used  to  supplement  the  coke  fed 
with  the  charge,  being  injected  at  the  tuyeres  of  the  blast-furnace,  and 
burned  as  it  meets  the  glowing  charge.  It  is  being  used  not  only  for  cop- 
per but  for  silver-lead  furnaces. 

In  smelting  ores  high  in  zinc,  this,  liberated  by  the  fuel  and  flux,  soon 
coats  the  descending  coke  with  a  white  coating  of  zinc  oxide,  so  that  it 
burns  with  difficulty;  the  temperature  falls,  the  slag  becomes  pasty  and 
works  poorly,  and  accretions  form,  so  it  is  evident  that  the  more  fuel  that 
can  be  added  at  the  tuyeres  the  better  the  furnace  should  run.  It  also  sug- 
gests the  smelting  of  such  ores  in  the  reverberatory  rather  than  in  the  blast- 
furnace. 

CALCULATION  OF  A  LEAD  BLAST-FURNACE  CHARGE 

When  sulphur-bearing,  oxidized,  or  silicious  ore  is  used,  we  have  to 
consider  not  only  the  sulphur,  silica,  and  other  constituents  of  the  ore, 
but  also  the  products  of  the  furnace  that  remove  the  constituents. 

Ore  (galena  for  example)  containing  less  than  10  to  12  per  cent  sulphur 
generally  is  smelted  without  roasting.  It  is  cheaper  to  do  this,  for  by  roast- 
ing, the  sulphur  of  the  ore  is  reduced  to  but  3  to  4  per  cent.  Many  ores 
within  the  above  limit  are  leady  ores,  and  difficult  to  roast  because  of  the 


CALCULATION  OF  CHARGE  485 

fusible  nature,-  but  the  matte  that  they  produce  is  easy  to  roast  for  the 
elimination  of  sulphur. 

Ore  intended  for  roasting  may  be  simple,  consisting  of  iron  sulphide,  or 
complex  as  shown  by  the  following  analysis  of  a  roasted  ore:  SiO2,  10 
per  cent;  Fe  and  Mn,  27  per  cent;  CaO,  Mgo,  and  BaO,  2  per  cent;  Zn, 
8.8  per  cent;  Cu,  0.4  per  cent;  S,  6  per  cent;  Pb,  35  per  cent,  and  Ag,  50 
oz.  per  ton.  The  base  in  the  roasted  ore  was  present  as  sulphide  in  the 
raw  ore. 

The  so-called  oxidized  ores  consist  of  the  carbonate  of  lead  with  a 
gangue  of  iron  oxide,  limestone,  dolomite,  and  silica.  Such  ores  though 
called  oxidized,  often  contain  a  little  sulphur,  as  sulphide  (galena  or 
pyrite)  or  as  sulphate. 

Silicious  ores  are  added  to  charges,  in  spite  of  the  large  excess  of  silica, 
because  the  gold  and  silver  are  present  in  quantity  to  pay  to  recover.  The 
lead  of  the  charge  takes  the  gold  and  silver  contained  in  such  ore,  while 
the  silicious  gangue  is  fluxed  into  a  barren  slag  and  sent  to  waste. 

Both  iron  ore  and  limestone  are  added  to  the  charge  for  fluxing  the 
silica,  making  a  slag  of  a  predetermined  composition  or  type.  If  the 
fluxes  contain  gold  or  silver  the  metals  can  be  recovered,-  since  they  go  into 
the  base-bullion  or  work-lead.  Without  gold  or  silver  they  are  called 
barren  or  "  dead  "  fluxes.  Ore  carrying  an  excess  of  iron  or  lime  over 
silica  (called  iron  or  lime  excess)  is  in  the  same  category,  since  the  excess  is 
useful  for  fluxing,  and  is  credited  in  purchasing  ores.  Thus,  ores  con- 
taining 10  per  cent  SiO2  and  40  per  cent  Fe  are  said  to  carry  30  per  cent 
iron  excess. 

Not  all  the  slag  that  issues  from  the  furnace  is  clean.  At  the  spout 
where  the  matte  flows,  and  in  the  shell  lining  the  cavity  of  the  fore-hearth, 
slag,  containing  drops  of  lead  and  matte,  is  found.  When  a  slag-pot  is 
emptied  at  the  edge  of  the  dump,  there  remains  a  shell  or  coating  of  soldi- 
fled  slag.  This  shell,  half  an  inch  thick,  is  found  to  contain  drops  of  matte 
that  did  not  entirely  settle  in  the  fore-hearth.  This  is  particularly  true 
when  the  fore-hearth  has  formed  a  thick  lining  and  soon  must  be  replaced 
by  another.  All  this  slag  having  value,  and  called  "  foul  slag,"  is  an  accept- 
able addition  to  the  charge  because  of  the  fusibility,  and  the  coarse  condi- 
tion, permitting  free  passage  of  the  air  of  the  blast. 

Computation  of  the  Charge. — To  determine  the  amount  of  the  fluxes 
(iron  ore  and  limestone)  to  add  to  the  charge,  to  give  a  slag  of  a  desired 
composition,  it  is  necessary  to  know  the  weight  of  the  ores  to  be  used,  and 
the  results  of  analysis  of  the  ore,  fluxes  and  fuel,  and  also  the  composition 
of  the  slag  and  matte  that  are  to  be  produced. 

When  a  charge,  thus  calculated,  has  been  put  on  the  furnace  and, 
after  several  hours,  has  come  down,  that  is,  has  come  into  the  melting  zone, 
so  that  the  slag  of  it  begins  to  flow  from  the  furnace,  then  a  sample  can  be 


486 


SILVER-LEAD  SMELTING 


taken,  analyzed  and  the  result  of  the  analysis  known  in  two  or  three  hours. 
If  this  result  shows  variation  from  the  desired  slag,  the  charge-composition 
may  be  suitably  altered  to  correct  it.  However,  before  making  any 
changes,  we  must  note  that  the  slag  is  hot  and  well  reduced. 

For  a  charge  to  contain  sintered  ore,  of  size  suited  to  automatic  or 
mechanical  charging,  the  following  is  an  excellent  example :  Of  this  charge 
some  70  per  cent  is  sintered  material  of  lead.  To  this  has  been  added  of 
silicious  ores,  A  and  B  350  Ib.  Bag-house  fume  is  a  product  burned  at 


BLAST-FURNACE  CHARGE  NO.  2 


WEIGHT. 

Pb. 

Si02. 

Fe(Mn). 

CaO. 

S. 

Wet. 

Dry.' 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 
Cent. 

Lb. 

Per 

Cent. 

Lb. 

Sinter  
Silicious  ore  A  
Silicious  ore  B  

3450 
250 
100 
100 
200 
800 

'700 
500 

24.0 
2.0 
2.5 
58.0 
3.0 

828 
7 
3 
58 
6 

20.5 
55.0 
50.0 
1.0 
11.0 
2.8 
5.0 

707 
137 
50 
1 
22 
22 
35 

25.0 
14.0 
16.0 
1.0 
41.0 
0.5 
2.0 

862 
49 
16 
1 
82 
4 
14 

5.4 

2.0 
9.0 
51.0 
0.5 

117 

'    2 
18 
408 
3 

4.4 

13.0 
6.2 

1.0 

151 

13 
6 

7 

Bag-house  fume  
Iron  ore  

Limestone  
Coke 

Slag  (shells,  etc.)  
Weight  of  charge.  .  .  . 

4900 

902 

974 

1028 

548 

177 

SiO2slag=    974 
FeO  +MnO  in  slag  =  1 107 


Fe  for  matte     168 

7)860  Sirfslag 


=40 


CaO  in  slag    =    548 

2629  =84  per  cent  of  slag 


123  S  volatilized  =53  93 

9  S  in  matte      =        84 

2 

1107  — 


Fe  for  matte  =      168 
Slag  — SiOz  =31  per  cent  Total  slag  =3130  Ib. 

FeO+MnO=35.5  84  380 

CaO=17.5  ^=380  matte;    —  =7.7  matte  fall. 

Matte,  Pb  =15.8  per  cent;   Fe  =45  per  cent;   Cu,  7.8  per  cent;   S  =22.5  per  cent;   Zn  =6.0  per  cent. 


the  bag-house  into  a  coherent  product.  To  these  is  added  iron  ore  and 
limestone  to  form  a  slag  to  contain  SiC>2,  31  per  cent;  FeO  (and  MnO); 
35.5  per  cent;  CaO,  17.5  per  cent.  There  is  8  per  cent  or  so  of  zinc  on 
the  charge,  and  the  slag  composition  corresponds  most  closely  to  the 
recalculated,  zinc-bearing  charge,  given  under  "  zinc."  The  slag  used  is 
higher  in  silica,  and  not  so  clean  as  the  type-slag;  on  the  other  hand  it 
runs  with  less  furnace  trouble.  Not  counting  the  coke  or  the  slag,  the 
charge  is  4900  Ib.  in  which  the  fuel  is  14.3  per  cent.  The  slag  (3130  Ib.) 
carries  1.5  per  cent  or  40  Ib.  of  sulphur.  Of  the  sulphur  30  per  cent  or 
53  Ib.  is  volatilized,  and  this  leaves  84  Ib.  to  form  matte.  It  will  be  seen 
that  the  ratio  of  sulphur  to  iron  gives  a  factor  of  2,  so  that  168  Ib.  of  iron  is 
needed  for  the  matte,  leaving  860  Ib.  of  iron  (equivalent  to  1107  Ib.  of 
FeO)  to  enter  the  slag.  Since  the  matte  contains  22  per  cent  of  sulphur 
the  84  Ib.  present  in  it  should  give  us  380  Ib.,  equal  to  a  matte-fall  of  7.7 
per  cent  only. 


CHAPTER  XXXVII 
PRODUCTS  OF  THE  BLAST-FURNACE 

FLUE-DUST 

A  blast-furnace,  44  by  154  in.,  takes  at  least  6000  cu.  ft.  of  air 
per  minute  when  in  full  operation.  The  escaping  gases,  of  an  average 
temperature  of  150°  C.,  have  expanded  to  8000  cu.  ft.,  and  have  a 
velocity,  while  rising  through  the  charge,  of  from  5  to  10  ft.  per  second. 
When  with  a  closed-top  furnace  the  side  doors  are  opened,  additional  air 
is.drawn  in  and  particles  of  20-mesh  size  may  be  carried  into  the  down-take 
and  the  long  main  flue  leading  to  the  tall  stack  which  produces  the  draft. 
This  main  flue,  which  is  common  to  all  the  blast-furnaces  of  the  plant,  is  of 
large  cross-section,  for  the  purpose  of  settling  and  collecting  the  particles 
called  "  flue-dust."  The  collected  dust,  after  suitable  preparation  either 
by  sintering  or  by  briquetting  under  pressure,  can  be  returned  as  part  of 
the  blast-furnace  charge.  It  commonly  amounts  to  0.5  to  0.8  per  cent  by 
weight  of  the  charge,  but  with  a  sinter  charge  it  should  be  even  lower  than 
this.  These  figures  refer  to  what  is  caught  in  the  flue.  Where  a  bag- 
house  is  used,  its  saving  should  be  added. 

Some  of  the  lead  and  silver  and  much  of  the  sulphur,  zinc,  and  arsenic 
of  the  charge  is  volatilized.  This,  in  part,  adheres  to  the  cool  surface  of 
the  flue.  Eventually  it  flakes  off  and  falls  to  the  bottom,  and  is  there 
recovered.  Flue-dust  is  therefore  composed  of  (1)  dust  carried  along  by 
the  flue  and  (2)  lead  fume,  containing  the  other  volatile  metals,  condensed 
on  the  cool  surface  of  the  flue. 

When  a  bag-house  is  not  used,  not  all  the  material  is  recovered;  the 
finest  part  may  escape.  An  analysis  of  flue-dust  made  at  the  Pueblo 
Smelting  Works,  Colo.,  shows  PbO,  37.6  per  cent;  ZnO,  53  per  cent; 
Fe2O3,  25  per  cent;  A12C>3,  1.3  per  cent;  CaO  (from  the  limestone),  5.3 
percent;  SiO2,  8.9  per  cent;  S,  2.5  per  cent;  SO3,  1.6  per  cent;  H2O,  CO2 
and  C  (from  the  coke),  11.2  per  cent.  When  arsenic  is  present,  part  con- 
denses in  the  flue,  but  much  of  it  escapes,  and  is  thus  happily  got  rid  of. 

The  carrying  power  of  the  moving  gases  varies  as  the  square  of  the 
velocity,  or  directly  as  the  draft  pressure,  therefore  to  settle  out  as  much 
flue-dust  as  possible  the  main  flue  should  have  a  large  sectional  area,  or 
better  yet,  if  the  bag-house  is  not  used,  a  dust-chamber  should  be  provided. 

487 


488 


PRODUCTS  OF  THE  BLAST  FURNACE 


Flues  have  been  made  of  sheet-steel,  but  the  metal  corrodes  under 
the  action  of  the  sulphuric  acid  and  sulphates  that  are  in  the  flue-dust, 
so  that  they  last  about  ten  years;  brick  therefore  remains  the  favorite 
material  for  such  construction.  The  bottom  of  brick  flues  is  frequently 
a  series  of  steel  hoppers.  Since  these  are  continually  covered  by  the 
flue-dust,  they  are  protected  from  the  fumes,  and  last  a  long  time.  A  flue, 
rectangular  in  cross-section,  may  have  brick  walls,  and  top,  and  a  hopper 
bottom  set  at  such  a  height  as  to  leave  room  beneath  for  a  car  running  on 
a  track  at  the  ground  level.  The  car  can  be  set  under  any  hopper,  and  to 
avoid  escaping  dust,  the  contents  may  be  drawn  into  the  car  through  a 
canvas  sleeve  fitted  over  the  spout. 


Bag  Chamber  C 

12" Cotton  Bags, 
34'3"long 


THE  BAG-HOUSE 

The  Bag-house. — Fig.  265  is  a  transverse  section  of  a  bag-house  140 

ft.  long  by  24  ft.  wide,  as 
used  for  the  recovery  of 
flue-dust  and  lead  fume, 
filtering  the  blast-furnace 
gases  or  fumes  through  cot- 
ton or  woolen  "  bags  "  that 
leave  them  colorless.  It  is 
of  a  size  to  take  the  gases 
from  six  furnaces,  equal  to 
60,000  cu.  ft.  per  minute. 

From  the  main  furnace 
flue  the  gases  are  drawn 
through  a  12-ft  pipe  called 
"the  trail,"  by  a  No.  14 
suction-fan  which  delivers 
them  under  pressure 
through  a  7-foot  pipe  to 
flue  A  of  the  bag-house. 
Just  before  entering  the 
fan,  at  the  top  of  tie  flue, 
there  is  fed  in  a  small  and 
regular  supply  of  finely 
ground  quicklime  to  neu- 
tralize the  small  amount  of 
H^SOa  in  the  gases,  which, 

if  allowed  to  remain,  would  speedily  corrode  the  bags. 

The  bag-house,  140  ft.  long,  consists  of  a  chamber  "B"  and  a  bag 

chamber  "  A,"  and  is  divided  by  transverse  partitions  into  five  chambers 


10  x  12  Ah 
Cylinders 


FIG.  265. — Bag-house. 


BAG-HOUSE 


489 


Canvas  Bag 


or  bays.  Each  bay  has  its  own  inlet  and  its  own  exhaust  pipe  both  con- 
trolled by  disk-valves,  and  an  exhaust  pipe,  branching  into  the  pipe  X, 
which  leads  to  the  trail.  As  shown  by  the  arrows  the  gases  pass  by  the 
inlet  into  the  chamber  "  B  "  enter  the  bags,  inflating  them,  and  filtering 
through  the  interstices  of  the  fabric,  escape  through  the  flue  "  D  "  to  the 
main  stack  (not  shown)  210  ft.  high.  In  the  course  of  eight  hours  the  bags 
are  getting  clogged  by  adherent  flue-dust  and  this  should  be  shaken  off. 
By  adjusting  the  two  disk  valves,  above  mentioned,  any  bag  is  by-passed, 
the  bags  of  it  are  collapsed  and  the  flue  dust  that  has  accumulated  on  them 
is  loosened  and  falls  down  into  the  chamber  "  B  "  of  that  bay.  The  bags  are 
shaken  several  times  at  five-minute  intervals  to  give  the  dust  time  to  fall 
down.  Once  in  several  days  a  side  door  of  the  bay  is  opened  and  a  man  en- 
tering takes  hold  of  and  gives  the  bags  a  more  thorough  shaking,  the  better 
to  clean  them.  The  dust  sliding 
down  the  steeply  inclined  bottom 
of  the  bay  is  carried  by  a  helical 
screw  as  in  Fig.  267  to  a  discharge 
opening  under  the  division  wall 
of  the  bay.  The  "  thimble  floor  " 
of  the  bag  chamber  "  C  "  is  of 
J-in.  sheet  steel,  pierced  with 
holes  for  thimbles,  Fig.  266, 
these  being  11  in.  in  diameter  by 

10  in.  high,  there  being  240  per       FIG  266._Detail  of  Bag-house  Thimble, 
bay,  or  1200  in  all.     The  bags  are 

about  13  in.  in  diameter  and  are  42  ft.  long,  wired  at  the  bottom  to  the 
thimbles  and,  by  means  of  wires,  closed  and  suspended  from  2-in.  pipes 
at  the  top. 

Treatment  of  the  Flue-dust. — The  dust  contains  60  per  cent  of  lead, 
and  is  extremely  fine.  When  2  ft.  in  depth  or  so  it  is  ignited,  and  once 
started,  combustion  proceeds,  causing  the  dust  to  become  sintered  together, 
and  in  a  condition  favorable  for  feeding  into  the  blast-furnace.  The  burned 
dust  contains  oxides  and  sulphates  of  the  metals,  viz.,  Pb,  60  to  70  per  cent; 
Zn,  3  per  cent;  Fe,  0.5  per  cent;  As,  1.3  per  cent,  and  Ag,  4.02  oz.  per  ton. 
Since  the  flue-dust  is  resmelted  and  the  arsenic  volatilizes,  it  tends  to  in- 
crease in  the  bag-house  product;  bismuth  accumulates  in  the  same  way. 


BRIQUETTING  FLUE-DUST 

Flue-dust  can  be  wet  down  and  fed  back  to  the  blast-furnace.  If 
fed  a  little  at  a  time,  it  is  simply  carried  again  into  the  flue,  but  while  wet, 
in  occasional  large  charges,  it  may  be  fed  so  that  most  of  it  is  carried  down 
and  smelted.  The  effective  way  is  to  make  it  into  briquettes  with  milk-of- 


490 


PRODUCTS  OF  THE  BLAST  FURNACE 


lime  as  a  binder.  Fig.  267  represents  a  plant  containing  a  White  briquet- 
ting-press  for  making  briquettes  composed  of  flue-dust  and  milk-of-lime 
to  which  is  added  fine  roasted  ore.  At  the  right  in  the  figure,  shown  to  be 
on  a  high  platform,  is  a  pile  of  quicklime.  This  is  fed,  together  with  water, 
into  the  lime-mixer,  a  trough  divided  transversely  by  a  partition.  One 
compartment  is  shown  as  containing  the  lime  being  mixed  to  a  thin  paste, 
while  the  other  is  now  empty.  The  paste  is  drawn  from  either  compart- 


FIG.  .267— White  Briquetting  Press. 


FlQ.  268.— Horizontal  Pug-mill. 

ment  to  a  horizontal  double-shaft  pug-mill.  Each  shaft  is  provided  with 
mixing-blades.  Flue-dust  from  the  pile  at  the  front  of  the  lower  pftitform 
is  shoveled  into  the  pug-mill  and  thoroughly  mixed  with  the  milk-of-lime 
by  the  revolving  blades  of  the  pug-mill  which,  being  set  at  an  angle,  propel 
it  to  the  discharge-opening  immediately  over  a  troughed  conveying-belt. 
It  drops  into  the  hopper  of  a  six-mold  briquetting-press  where  it  is  made 
into  briquettes  that  drop  upon  a  flat  conveying-belt  delivering  them  to  a 
pile.  The  briquettes  may  be  used  at  the  blast-furnace  freshly  made,  but 
the  usual  plan  is  to  dry  them,  as  clay-bricks  are  dried. 


METHODS  OF  MATTE  TREATMENT  491 

Fig.  268  illustrates  the  internal  construction,  the  spiral  screw  at 
one  end  of  the  shaft  acting  to  speedily  discharge  the  mixed  material. 

At  times  the  briquetting  is  omitted,  and  the  pug-mill  mixture  is  wheeled 
to  a  drying-floor,  or  is  distributed  evenly  upon  one  of  the  ore-beds.  By 
the  time  the  bed  is  used  the  mixture  has  set  and  becomes  a  hard  mass 
capable  of  withstanding  handling  without  being  broken. 

LEAD-COPPER  MATTE 

The  lead-smelting  charge  generally  contains  copper,  and  the  copper 
accumulates  in  the  matte.  Since  matte  is  roasted  and  returned  to  the 
blast-furnace,  the  content  in  copper  gradually  increases.  When  increased 
to  12  per  cent,  the  copper  matte  is  again  roasted  and  treated  in  a  separate 
blast-furnace,  with  silicious  ore  and  oxidized  copper  ore,  to  produce  a 
matte  of  40  per  cent  copper,  called  "  shipping-matte  "  because  it  often  is 
shipped  to  a  copper  works  to  be  treated  for  copper.  This  operation  is 
called  "  concentrating." 

COMPARISON  OF  MATTE-TREATMENT  METHODS 

It  has  been  urged  against  the  treatment  of  low-grade  matte  that  from 
each  ton  treated  in  concentrating  it  there  is  produced  a  ton  of  slag,  and 
half  a  ton  of  lead-zinc  fume  both  of  which  have  to  be  retreated  in  the 
blast-furnace.  But  in  the  case  of  shipping  matte,  3  tons  of  low-grade 
matte  having  been  made  into  1  of  shipping  grade,  the  consequent  slag 
per  ton  of  matte  is  but  half  a  ton  and  the  percentage  of  lead  is  propor- 
tionately lower. 

The  average  composition  of  shipping  matte  may  be  thus  given: 

Pb,  26.9  per  cent;  Cu,  43.1  per  cent;  Ni  and  Co,  0.4  per  cent;  Si(>2,  0.3 
per  cent;  Fe,  8  per  cent;  Zn,  2.5  per  cent;  S,  15.5  per  cent;  As,  1.7  per 
cent;  Sb,  0.76  per  cent;  showing  the  complex  nature  of  such  matte,  and 
how  it  takes  up  every  impurity. 

A  satisfactory  way  of  treating  matte,  where  it  is  wished  to  produce  a 
more  finished  product,  is  to  crush  it  to  a  4-mesh  size  and  to  roast  it.  It  is 
next  sent  to  a  blast-furnace  and  again  smelted  with  silicious  and  oxidized 
ores  of  high  grade  in  copper.  There  results  a  matte  of  65  per  cent  Cu  and  a 
certain  quantity  of  "  bottoms,"  the  result  of  the  separation  of  copper  from 
the  matte.  The  bottoms  are  charged  into  a  reverberatory  furnace  through 
side-doors,  and  the  coarse-broken  matte  is  put  on  top.  The  doors  are 
closed,  and  the  charge  is  fired  with  an  oxidizing  flame,  as  in  the  Welsh 
process  of  "  roasting."  The  charge  having  melted,  a  reaction  of  the 
cuprous  oxide  on  the  cuprous  sulphide  takes  place,  as  described  in  the 
Welsh  process  of  making  blister-copper,  see  page  389,  and  the  charge 


492  PRODUCTS  OF  THE  BLAST  FURNACE 

becomes  reduced  to  an  impure  copper  containing  arsenic,  bismuth,  and 
antimony,  as  well  as  the  gold  and  silver  that  were  contained  in  the  matte. 
The  copper  is  then  poled  to  reduce  the  cuprous  oxide,  and  ladled  into  anode- 
molds.  The  anodes  are  sent  to  an  electrolytic  copper  refinery  for  treat- 
ment. 

THE  CONVERTING  OF  LEADY  MATTE 

One  smelting  works  in  the  United  States,  the  Tooele  plant  of  the 
International  S.  &  R.  Co.,  has  one  department  for  the  smelting  of  silver 
lead  ores,  another  for  the  reverberatory  smelting  and  converting  of  copper 
ores.  It  is  therefore  easily  possible  to  treat  the  matte  from  the  silver- 
lead  furnaces  in  the  converter  as  follows: 

The  matte  is  tapped  from  the  blast-furnace  fore-hearth  into  10-ton  pots 
and  at  the  converter  department  is  poured  into  similar  pots  standing  in  a 
pit.  The  final  portion  is  held  back  to  be  poured  into  a  shallow  cast-iron 
pan  together  with  some  lead  that  has  separated  out  in  the  bottom  of  the 
pan.  When  cold  enough  the  crust  of  matte  is  lifted  off,  leaving  the  lead, 
still  molten,  behind.  The  main  body  of  the  matte  in  the  10-ton  pot  is 
transferred  by  crane  and  poured  into  the  converter.  At  the  beginning  of 
the  blow  copious  fumes  are  evolved  which  are  drawn  by  a  suction-fan  to  a 
bag-house.  Special  care  must  be  taken  that  these  fumes  do  not  get  too 
hot  so  as  to  set  the  bags  afire.  Pyrometers  connected  to  signal  lights  indi- 
cate when  this  may  happen  and  the  converter  man  may  then  turn  off  the 
blast  until  the  temperature  becomes  normal  again.  The  quantity  of  fume, 
of  both  lead  and  zinc,  diminishes  as  the  blow  proceeds,  the  smaller  quan- 
tity of  zinc  fume  uniting  itself  to  any  SOs  present  to  form  zinc  sulphate 
and  so  preventing  acid  corrosion  of  the  bags.  The  bulk  of  the  fume  is 
lead,  being  about  65  per  cent  of  the  total.  This  fume,  collecting  at  the 
bottom  of  the  lower  chamber  of  the  bag-house  is  set  afire  from  time  to 
tune,  forming  a  slightly  sintered  product  to  return  to  the  blast-furnace. 

The  converter  is  blown  without  addition  of  silica,  thus  forming  a  slag 
high  in  magnetite,  this  magnetite  forming  a  basic  lining  that  is  permanent. 
The  slag  itself  is  basic.  The  contents  of  the  ladle  go  to  a  granulator, 
where  it  is  poured,  the  stream  of  slag  being  hit  by  a  flat  jet  of  water  which 
granulates  it.  The  product  sinks  into  a  deep  hopper  filled  with  water,  out 
of  which  the  slag  is  raised  by  a  bucke1>-elevator.  Added  to  the*roaster 
charge,  it  forms  an  acceptable  item  of  the  sinter  charge  because  of  its  excess 
of  iron. 

In  a  typical  instance  (per  ton  of  blister  copper  produced),  there  is 
yielded  in  the  flues  and  at  the  bag-house  0.368  ton  of  fume,  yielding 
0.221  ton  of  lead.  An  analysis  of  the  fume  showed  Pb,  52.5  per  cent; 
Cu  trace;  Zn,  3  per  cent;  S,  5.4  per  cent;  As,  14.2  per  cent;  Sb,  1.6  per 
cent;  Fe,  trace;  Ag,  10  oz.  per  ton.  This  indicates  that  copper  is  not 


PRICE  OF  MATTE  493 

volatile,  while  arsenic  is  quite  so.     The  fume,  after  burning,  is  sent  to  the 
blast-furnace  to  recover  the  lead. 

Returning  to  the  matte,  now  freed  from  lead  and  zinc,  it  is  added  to  the 
copper  converter  charge,  the  whole  being  blown  to  blister  copper. 

SELLING  PRICE  OF  MATTE 

Gold  is  paid  for  at  $19  per  ounce,  silver  at  95  per  cent  of  the  New 
York  quotation,  copper  price  to  be  that  of  electrolytic  at  New  York  quo- 
tation. Lead  over  5  per  cent  is  charged  for  at  30  cents  per  unit,  zinc  over 
7  per  cent  at  20  cents  per  unit. 


CHAPTER  XXXVIII 


PRODUCTION  OF  LEAD  ORES  AND  PRICES 

COSTS  OF  LEAD  ORES 

To  illustrate  the  method  of  calculating  the  actual  cost  of  treating  an 
ore,  as  in  the  Colorado  or  Utah  silver-lead  smelting  practice,  we  take  the 
case  of  a  so-called  neutral  ore  (SiCb  equal  to  Fe) .  The  ore  is  assumed  to  be 
oxidized,  to  contain  less  than  5  per  cent  sulphur,  and  at  least  10  per  cent 
lead.  It  is  to  be  treated  at  a  works  having  an  output  of  400  tons  of  charge 
daily. 

The  cost  of  treating  a  ton  of  charge,  and  of  treating  a  ton  of  the  ore 
including  the  flux,  is  as  follows: 


Costs  per  Ton. 

Charge. 

Ore. 

Labor  

1.76 

2.46 

Coke  (12  6  per  cent  of  charge  at  $12  60  per  ton) 

1  60 

2.24 

Power  costs                                                      

0.14 

0.19 

Interest  replacements  and  repairs 

0.52 

0.73 

Delays  due  to  accidents  strikes  etc    5  per  cent 

0  37 

0  52 

General  expense  assaying  and  management  

0.40 

0.56 

Depreciation  10  per  cent  on  $2,000,000  investment  of  plant.  .  .  . 
Limestone  (0  3  ton  at  $2) 

0.04 

0.06 
0.60 

Iron  ore  (0  1  ton  at  $8)                           

0.80 

$4.83 

$8.76 

The  figures  in  the  second  column  are  obtained  by  multiplying  the  total 
weight  of  the  charge,  1.4  tons,  by  the  cost  of  each  item  per  ton  of  material 
and  then  adding  the  cost  of  the  flux.  This  corresponds  to  the  figure  above 
obtained,  and  may  be  stated  again  as  follows : 

1.4  tons  of  material  (1  ton  ore,  0.3  ton  limestone,  and  0.1  ton  of  iron  ore)  at  $4.83 

per  ton  of  charge  for  smelting $6 . 86 

Cost  of  fluxes,  0.3  ton  of  limestone  at  $2.00 0.60 

Cost  of  fluxes,  0.1  ton  of  iron  ore  at  $8.00 0.80 


$8.16 

In  case  the  ore  contains  sulphur  in  quantity  to  require  roasting,  $2 
should  be  added.     For  sulphur  over  5  per  cent  and  up  to  10  per  cent  add 

494 


ORE  PRICES  495 

30  cents  per  unit  to  cover  the  expense  of  iron  ore  for  disposing  of  the  extra 
sulphur,  and  roasting  the  matte  made  by  it. 

Distribution. — The  costs  of  production  per  ton  of  charge  smelted 
have  been  thus  divided:  Labor  23  per  cent;  coke,  40  per  cent;  coal,  5 
per  cent;  limestone  for  coke  ash,  5  per  cent;  maintenance  and  repairs, 
5  per  cent;  delays  due  to  accidents,  strikes,  etc.,  5  per  cent;  flue  dust 
recovery,  2  per  cent;  administration,  7  per  cent. 

ORE  PRICES;    MISSISSIPPI  VALLEY  LEAD  SMELTING  WORKS 

Non-argentiferous  lead  concentrates  are  bought  at  a  quoted  rate  based 
upon  an  80  per  cent  lead  content  as  determined  by  wet  assay,  with  a 
deduction  of  50  cents  per  unit  for  all  below  and  an  addition  of  50  cents  for 
all  over  80  per  cent. 

ORE  PRICES;  COLORADO  AND  UTAH  SILVER-LEAD  SMELTERIES 

Silver-lead  and  Dry  Ores. — The  price  per  ton,  dry  weight,  delivered  at 
the  smelting  works,  depends  upon  the  value  of  the  contents  of  the  ore  as 
determined  by  fire  assay,  and  based  upon  the  New  York  price  of  the 
metals.  From  these  values  must  be  deducted  the  charge  for  treatment  or 
working  charge  and  a  reasonable  profit.  The  working  charge  (W.  C.) 
varies  with  the  lead  contents,  the  insoluble  residue  (approximately  the 
silica),  the  iron,  the  sulphur,  zinc,  and  speiss  (iron  arsenide)  present 
in  the  ore. 

A  central  custom  silver-lead  smeltery  buys  and  combines  for  smelting  a 
variety  of  ores  to  its  profit;  and  smelts,  not  only  argentiferous  lead  ores, 
but  lead-free  or  dry  ones.  This  it  can  do  if  it  has  enough  lead-bearing 
ore  on  the  charge  to  insure  the  extraction  of  the  precious  metals  from  the 
dry  ores  also  being  smelted. 

The  Metal  Values. — These  are  commonly  paid  for  as  follows:  Gold  at 
$19  per  ounce,  Silver  at  95  per  cent  of  its  New  York  value  less  a  further 
deduction  in  Utah  of  3.5  cents  per  ounce  to  cover  the  freight  charge  now 
made  upon  base-bullion  that  contains  it,  but  no  such  charge  is  made  upon 
silver  in  Colorado  ores.  The  5  per  cent  deduction  made  in  the  silver  price 
is  intended  to  cover  that  lost  in  smelting. 

Lead. — Based  upon  a  quotation  of  6  cents  per  pound  a  deduction  of 
10  per  cent  is  made  to  cover  the  smelting  loss  and  in  Utah  a  further  deduc- 
tion of  $35  per  ton  of  lead  to  cover  freight  and  refining  loss.  This  is  based 
upon  a  charge  of  $17  per  ton  of  lead  for  freight  and  a  further  deduction 
of  $18  per  ton  for  refining. 

Treatment  Rate  or  Working  Charge. — The  base-price  for  smelting  ores 
is  $2.50  per  ton  with  lead  at  6  cents  per  pound;  if  the  value  of  lead  exceeds 
this,  then  50  cents  per  unit  is  added  to  the  treatment  rate,  while,  when 


496  PRODUCTION  OF  LEAD  ORES  AND  PRICES 

lead  falls  below  6  cents,  a  corresponding  50  cents  a  unit  is  taken  from  that 
rate.  No  distinction  is  now  made  between  oxidized  and  sulphide  ores. 
At  present  the  furnace  charge  is  two-thirds  or  more  sinter,  and  in  this  has 
been  put  the  fines  out  of  all  ores,  such  fines  coming  from  the  custom  of 
crushing  when  sampling.  Since  the  cost  per  ton  of  ore  furnaced,  and  the 
cost  per  ton  of  materials  put  through  is  accurately  known,  as  well  as  the 
cost  of  roasting  and  refining,  it  has  been  thought  best  to  make  the  base 
price  for  smelting  equal  to  this  actual  cost  plus  a  reasonable  profit  and  to 
seldom  make  finer  distinctions.  This  runs  much  the  same  day  by  day,  and 
the  month's  profit,  in  normal  operation,  approaches  the  known  figure. 
When  lead  ores  are  scarce  the  ore  buyer  will  make  concessions  to  procure 
them.  In  Colorado  a  different  rule  holds  and  in  the  schedule  for  dry 
oxidized  ores  the  treatment  charges  vary  from  $7.50  to  $9.50  per  ton  be- 
tween gross  values  of  from  $8  to  $20,  that  is  these  charges  increase  with 
the  value  of  the  ore. 

Debits  or  Penalties. — A  charge  is  made  of  10  cents  per  unit  for  all 
insoluble  matter.  Speiss  over  5  per  cent  is  charged  at  20  cents  per  unit. 
All  zinc  over  10  per  cent  is  charged  at  30  cents  per  unit.  All  sulphur  is 
charged  at  25  cents  a  unit,  but  not  to  exceed  a  maximum  of  $3  per  ton. 

PROFITS  PER  TON 

We  will  now  compute  the  price  f .  o.  b.  at  the  works  for  a  Utah  ore  of  the 
composition,  Insol.,  40  per  cent;  Fe,  10  per  cent;  CaO,  5  per  cent;  Pb, 
20  per  cent;  Cu,  0.3  per  cent;  Zn,  8  per  cent;  S,  5  per  cent;  Ag,  18  oz. 
and  Au  0.10  oz.  per  ton. 

Credits. 

Metal  values  Au  0.10  oz.  at  $19 $1 .90 

Ag  18  oz.  at  $1.10  per  oz.  95  per  cent  of  this  less  3£  cents 18 . 18 

Pb  20  per  cent  or  400  Ib.  less  10  per  cent  less  1.75  cents  with  lead  at 

at  6  cents  per  pound 15 . 30 


Metal  values $35.38 

Fe  credit  10  per  cent  at  6  cents 60 


Total  credits $34 . 78 

Debits.  t 

Treatment.     Base  charge $2 . 50 

Insoluble  40  per  cent  at  10  cents 4 . 00 

Zinc  5  per  cent  over  allowance  at  30  cents 1 . 50 


Total  debits. .  .$8.00          $8.00 


Net  value  of  the  dry  ore $26 . 78 

The  realization  per  ton  will  be 


SMELTING  COSTS  497 

Gold,  100  per  cent  of  0.10  oz.  at  $20.56  per  ounce $2 . 06 

Silver,  98  per  cent  of  18  oz.  at  $1.10  per  ounce 19 .40 

Lead,  92  per  cent  of  400  Ib.  at  6  cents  per  pound 22 . 08 


Total  metal  values • $43 .54 

Net  cost  per  ton  f  .o.b.  at  the  works 26 . 78 

Treatment 6.00 

Freight  at  $17,  refining  at  $12  on  0.184  ton  base  bullion 5 . 34 

Selling  the  product. .  . 0 . 32 

Interest  on  metals  in  process,  three  months  at  6  per  cent 0.53 


Total  costs $38.97 

Profits  per  ton  to  pay  for  capital  investment  to  balance 4.57 

Total  realization ..V;  * $43.54 

Since  analysis  is  based  on  the  dry  ton  the  moisture  must  be  deducted  from 
the  gross  weight  to  determine  the  weight  dry.  No  charge  is  made,  for 
sampling  except  in  lots  of  a  few  tons. 

VARIATION  IN  COSTS  DUE  TO  OUTPUT,  ETC. 

Comments  on  Costs. — Where  operations  proceed  smoothly  where  slag 
losses  are  no  greater  than  given  for  good  smelting,  where  by  the  bag-house 
and  the  electrostatic  treater  the  metal  losses  are  low,  where  gold  and  cop- 
per in  small  amounts  are  not  paid  for,  and  where  the  metals  are  skillfully 
sold,  profits  may  be  as  high  as  above  calculated.  On  the  other  hand,  if  iron 
ore  and  limestone  must  be  purchased  for  fluxing,  if  the  metal  market  is  a 
falling  one,  if  it  is  not  possible  to  get  the  best  combination  of  ores  for 
profitable  smelting,  if  the  works  are  running  at  part  capacity,  these  profits 
will  dwindle.  The  money  to  carry  a  stock  of  ores  -and  supplies,  and  for 
freight  advances  should  be  valued  at  6  per  cent  per  annum. 

To  treat  a  silicious  ore  of  50  per  cent  silica,  provided  the  iron  ore  and 
limestone  must  be  purchased,  is  expensive,  since  with  1  ton  of  ore  one  must 
use  1.5  tons  of  fluxes  and  smelt  2.5  tons  of  materials  of  the  charge,  the 
estimated  cost  being  $21  per  ton.  The  actual  charge  for  treatment  is 
about  half  this  and  even  the  extra  charge  of  3J  cents  per  ounce  of  the  silver 
does  not  compensate.  In  general  there  is  an  excess  of  iron  for  fluxing  which 
comes  from  iron  sulphides  and  irony  ores.  It  must  also  be  remembered 
that  a  ton  of  such  silicious  ore  will  produce  1.75  tons  of  slag,  which  slag 
carries  off  silver  and  lead.  The  term  "  displacement  "  refers  to  that 
condition  where  much  silicious  ore  is  smelted,  so  that  for  100  tons  smelted 
40  tons  would  be  ore,  while  if  the  ore  is  neutral  of  100  tons  of  charge  72.5 
tons  would  carry  the  profit. 

The  price  may  be  modified  according  to  the  needs  of  the  works.  Thus, 
if  lead  ores  are  much  needed  they  may  be  bought  even  at  a  loss  while  to 
compensate  the  more  plentiful  ores  may  be  bought  at  a  low  price. 


CHAPTER  XXXIX 
REFINING  OF  LEAD  AND  BASE-BULLION 

Primary  lead,  or  that  produced  by  smelteries,  may  be  divided  into 
three  kinds  on  the  market,  viz. :  Soft  lead,  which  comes  from  non-argen- 
tiferous ores;  desilverized  lead  produced  from  base-bullion;  antimonial 
lead,  a  by-product  of  the  Parkes  process.  Soft  lead  from  the  ore-hearth 
is  commonly  remelted  and  poled  to  remove  impurities.  Desilverized  lead 
may  be  divided  into  common  lead  suited  to  making  pipe,  sheet  lead,  shot, 
and  lead  alloys.  A  softer  grade  is  called  corroding  lead  for  making  white 
lead.  Antimonial  lead  as  a  base  for  type  metal  and  bearing-metal  con- 
tains 15  to  20  per  cent  antimony. 

REFINING  BASE-BULLION 

Sampling  and  Handling. — The  practice  at  large  silver-lead  smelting 
works  is  now  to  remelt  all  base-bullion  from  the  blast-furnace.  Some- 
times the  lead  is  taken  in  molten  condition  to  the  remelting  kettle  from 
the  blast-furnace.  When  melted  in  the  remelting  kettle  it  is  carefully 
skimmed,  and  as  in  lead  refining,  the  cleaned  lead  is  molded  into  bars. 
The  skimming  or  dross,  containing  copper  and  other  impurity,  is  returned 
to  the  blast-furnace.  The  copper  there  enters  the  matte  and  the  lead 
again  goes  to  the  base-bullion.  While  the  lead  is  being  molded  samples 
are  taken  from  the  kettle  at  intervals,  and  from  the  samples  the  assay- 
results  are  obtained.  The  bars  for  a  40-ton  carload,  800  in  number,  are 
stamped  with  the  number  of  the  lot,  and  are  carefully  weighed,  twenty  at 
a  time.  Careful  assays  are  made  of  each  lot,  both  by  the  shipper  and  by 
the  refiner. 

At  smaller  plants  the  punch  sample  is  taken  as  described  in  the  chapter 
on  sampling.  The  results  are  exact. 

Characteristics  of  Base-bullion. — Lead  containing  silver,  commonly 
called  base-bullion,  is  refined  by  the  Pattinson  or  by  the  Parkes  process. 
Commonly  the  Parkes  process  is  used.  The  object  in  either  process  is  to 
effect  the  separation  of  the  silver  and  gold  from  the  lead. 

To  get  a  clear  idea  of  the  principles  of  refining  base-bullion  (or  work- 
lead  as  it  is  called  in  Europe)  we  first  must  know  the  composition.  An 
especially  base  quality  is  represented  by  the  following  analysis : 

498 


THE  LEAD  REFINERY 


Per  Cent. 


Impurities*  Cu                                       

0.82 

As 

0.38 

Sb  

0.71 

Fe              

0.02 

8.. 

0.14 

499 


Per  Cent. 
96.59 


Precious  metals:  Ag  (322  oz.  per  ton) 1 . 07 

Au  (0.20  oz.  per  ton) 0.0007 


2.67 


1.07 


99.73 


FIG.  269. — Lead-refinery  Building  (cross-section). 

It  is  seen  that  base-bullion  is  principally  lead.  The  problem  is  to 
soften  the  lead  by  removing  the  impurities,  and  then  to  separate  the  gold 
and  silver  from  the  purified  or  softened  lead.  In  studying  the  process  the 
student  should  refer  to  Figs.  280  and  269. 


THE  REFINERY 

In  Fig.  269  we  have  a  cross-section  of  the  refinery  building  showing  the 
course  of  the  base-bullion  through  it  as  further  elucidated  in  the  flow-sheet 
Fig.  270. 

The  bullion,  in  bars  or  ingots  of  100  lb.,  pass  by  an  inclined  elevator 
to  the  softening  furnace  *S,  whence  the  softened  metal  goes  to  the  zincing 
kettle  K.  The  silver  and  gold  are  here  removed  and  the  lead  yet  con- 
taining some  zinc  passes  on  to  another  reverberatory  furnace,  where  the 


500 


REFINING  OF  LEAD  AND  BASE-BULLION 


zinc  is  burned  off  and  the  residual  lead,  now  soft,  is  molded  into  bars  or 
ingots  (see  Fig.  276),  for  the  market. 

Fig.  270  is  a  flow-sheet  of  the  process  from  the  receiving  of  the  base 
bullion  to  the  production  of  the  market  lead  and  silver-gold  or  dore  bars, 
the  latter  to  be  parted  for  the  production  of  gold  and  silver. 

SOFTENING  BASE-BULLION 

The  Softening  Furnace. — Softening  is  performed  in  a  water-jacketed 
reverberatory  furnace,  Fig.  271.  The  rectangular  hearth  of  the  furnace, 
7  by  14  ft.  in  size,  is  surrounded  by  a  sheet-steel  double  water-jacket, 
shown  in  section  at  (a)  in  the  sectional  elevation  (c).  The  jacket  assists 

Base  Bullion 


-1st  Skim,  to  Blast  Furnace 

-2nd     "       "  Precipitating  Furnace 

-3rd     «       "    Blast  Furnace 

Rrtort  Zinc 


•Lead— <- 


/Zin 
^U    Ke 

Desilveri 

-'"eV^  l8t  Crust 

. 

(.^.j^      1         2nd  Crust  to 

Retort 

y^\  next  chaw 

Rich  Lead 

;ed  Cfea95 

Cupelling 
Furnace 

Calcining 
Furnace 

Silvel  Bars 

Litharge  to 
Blast  Furnace 


Skim  to  Blast  Furnace        Skim  to  Blast  Furnace 

FIG.  270. — Lead-refining  by  the  Parkes  Process. 

in  resisting  the  action  of  the  molten  litharge  formed  from  the  lead  in  the 
operation.  Within  the  water-jackets  the  hearth  and  walls  are  a  dense 
aluminous  brick,  and  at  the  slag  line,  where  the  corrosive  action  of  the 
litharge  is  intense,  bauxite  brick  of  98  per  cent  A^Oa  are  used.  The  fur- 
nace is  heated  by  the  firebox,  having  a  grate  4  by  5  ft.  in  dimensions,  so  that 
a  high  temperature  can  be  attained  in  the  furnace.  The  letters  c,  c  indicate 
the  rear  working  doors  and  6,  6,  b  the  front  doors  in  which  the  base^bullion 
is  charged.  At  the  front  of  the  furnace  the  tap-hole  e  is  provided,  through 
which  the  lead  is  tapped  when  the  charge  is  finished.  The  furnace  com- 
municates to  a  stack  50  ft.  high  by  a  flue  at  the  front  end. 

Operation. — The  work  is  done  in  two  stages.  In  the  first  stage  at  a  low 
temperature,  the  copper  is  removed.  In  the  second,  at  a  high  temperature, 
the  arsenic  and  antimony  are  expelled,  after  which  there  is  left  only  the 
softened  lead  containing  the  precious  metals. 


SOFTENING  FURNACE 


501 


The  base-bullion,  in  charges  of  30  tons,  is  placed  in  the  furnace  by  means 
of  a  long-handled  paddle  or  "  peel  "  having  a  blade  2  ft.  long  by  6  in.  wide. 
The  bars  are  laid  one  at  a  time  upon  this  and  placed  as  desired  in  the  fur- 
nace, being  piled  in  a  heap  on  the  hearth.  The  doors  are  closed  and  the 
bars  are  gradually  melted  down,  the  dross  contained  in  the  bullion  rising 
to  the  top.  When  melted  the  heat  is  maintained  slightly  above  the  melt- 
ing-point, but  not  higher.  In  about  two  hours  the  dross  that  has  risen  to 
the  top  is  carefully  skimmed,  by  means  of  a  long-handled  perforated 


FIG.  271. — Softening  Furnace  (Parkes  process). 


skimmer,  and  removed  through  the  door  to  a  wheelbarrow  placed  for  it. 
The  dross,  residue,  or  skimming,  called  the  "  copper  skim/'  consists  of  a 
drossy  lead  containing  the  iron,  sulphur,  and  (especially  important)  most 
of  the  copper  of  the  base-bullion.  The  removal  of  these  completes  the 
first  stage  of  the  process.*  The  liquated  dross  thus  skimmed,  which 

*  It  should  be  here  noted  that  where  the  smelting  plant  and  refinery  are  in  one,  the 
blast-furnace  lead  is  treated  as  described  under  "  Sampling  and  Handling."  There 
results  a  product  which,  when  charged  into  the  softening  furnace,  can  omit  the  first 
stage  and  the  lead  is  treated  to  the  second  stage  or  obtaining  the  "  antimony  skim  " 
only. 


502 


REFINING  OF  LEAD  AND  BASE-BULLION 


may  amount  to  5  per  cent  of  the  charge  or  1.5  tons,  consists  of  Pb,  62.4 
per  cent;  Cu,  17.97  per  cent;  Ag,  0.17  per  cent  (49  oz.  per  ton);  As,  2.32 
per  cent;  Sb,  0.98  per  cent;  Fe,  0.43  per  cent;  S,  4  per  cent;  and  0,  1.87 
per  cent.  Slag,  ash,  and  hearth  material  also  are  contained  and  must  be 
reckoned  in. 

The  heat  of  the  molten  bath  is  now  raised  to  a  bright  red  (600°  to  650° 
C.)  and  the  flame,  made  oxidizing  by  the  admission  of  an  excess  of  air 
through  the  thin  fire,  sweeps  over  the  surface.  Litharge  forms,  and  the 
antimony  and  arsenic  oxidize  and  enter  the  litharge  slag.  The  litharge 


FIG.  272. — Howard  Mixer. 

at  this  temperature  has  a  corrosive  action  upon  the  brick  lining,  hence 
the  need  of  a  water- jacketed  furnace.  This  stage  of  the  process  lasts 
twelve  hours,  until  a  sample  of  the  lead  taken  from  the  furnace  and  placed 
in  a  mold  and  skimmed,  shows  by  the  appearance  that  it  is  free  from  arsenic 
and  antimony.  Before  the  antimony  is  removed  the  surface  of  the  molten 
lead  will  "  work,"  or  show  oily  drops  moving  upon  it.  A  similar  phe- 
nomenon is  seen  in  the  first  stage  of  cupelling  base-bullion  high  in  anti- 
mony and  arsenic.  As  the  softening  proceeds  the  drops  become  fewer  and 
smaller,  and  finally  a  coating  is  seen  to  dull  the  surface  of  the  hot  molten 
lead,  indicating  the  completion  of  the  softening.  For  impure  base-bullion 
this  stage  is  of  more  than  twelve  hours'  duration,  and  the  thick  layer  of 
litharge  formed  retards  further  oxidation.  It  is  best  then  to  draw  the  fire 


THE  PARKES  PROCESS  503 

and  to  cool  the  charge,  to  allow  the  litharge  slag  on  the  top  to  solidify  above 
the  liquid  lead  beneath.  The  slag  is  skimmed  with  a  long- handled  per- 
forated skimmer  (compare  with  Fig.  275),  and  the  charge  is  fired  again 
if  necessary  until  the  impurities  are  removed.  The  "  antimony  skim  " 
consists  of  the  antimonate  and  arsenate  of  the  lead  with  a  large  proportion 
of  litharge.  It  is  in  fact  an  impure  litharge  containing  15  to  20  per  cent 
antimony. 

The  softened  lead  to  be  treated  by  the  Pattinson  (see  page  510),  or  the 
Parkes  process  for  the  removal  of  the  contained  gold  and  silver,  is  now 
tapped  into  the  desilverizing  kettle,  8  ft.  diameter  and  capable  of  hold- 
ing 30  tons  of  lead. 

THE  PARKES  PROCESS 

Operation. — The  softened  lead  from  the  softening-furnace  is  tapped  into 
a  hemispherical  cast-iron  kettle,  shown  in  Fig.  272,  which  holds  30 
tons  or  more  of  lead  or  the  full  charge  from  the  softener.  The  kettle  is 
set  hi  brickwork,  and  is  heated  from  a  firebox  below.  In  modern  practice 
kettles  are  made  large  and  are  10  ft.  diameter  and  2  ft.  10  in.  deep,  holding 
60  to  65  tons. 

The  lead  from  the  softening-furnace  flows  along  a  cast-iron  trough  to 
the  kettle.  In  so  doing  a  litharge  dross,  called  "  kettle  dross,"  forms,  and 
collects  on  the  surface  of  the  metal,  and  is  skimmed  off. 

The  principle  of  the  separation  of  silver  from  lead  depends  on  the 
affinity  of  silver  for  zinc,  which  is  greater  than  for  lead.  Upon  adding  and 
thoroughly  mixing  in  a  small  amount  of  zinc  it  takes  up  most  of  the  silver. 
Zinc  has  a  greater  affinity  than  lead,  not  only  for  silver,  but  for  gold  and 
copper.  When  the  molten  bath  is  allowed  to  stand  a  while,  the  zinc,  being 
lighter,  separates  and  rises  to  the  sur- 
face. At  a  temperature  below  the 
melting-point  of  zinc,  but  above  that 
of  lead,  a  crust  forms  that  can  be 
skimmed  off.  Thus  the  silver  is  con- 
centrated in  a  small  bulk  of  metal, 
and  is  later  separated  from  the  rich 
metal  by  further  treatment. 

The  molten  bath  is  heated  to  an 
incipient    red    heat,    well   above   the 
melting-point   of   zinc,  and   cakes   or 
ingots  of   spelter  equal  to   about  1.2        FlQ  273.-Skirnming  Base-bullion, 
per   cent  of    the  weight,  or    720  lb., 

are  added.  The  quantity  required  varies  with  the  richness  of  the  base- 
bullion  in  silver.  The  added  zinc  quickly  melts. 

Desilverizing  machinery   is  now  much  used.     The   most   approved 


504 


REFINING  OF  LEAD  AND  BASE-BULLION 


machine  is  the  Howard,  used  both  for  mixing  and  skimming.  Fig.  272 
represents,  in  section  and  elevation,  the  kettle  and  the  apparatus  used  for 
intimately  mixing  the  molten  zinc  with  the  lead.  The  machine  is  brought 
to  the  kettle  by  an  overhead  crawl  h  and  is  lowered  into  it  by  a  chain-block 
hoist.  When  lowered  into  position  as  shown  in  Fig.  272,  the  screw  pro- 


FIG.  274. — Howard  Press. 


peller  b  is  set  in  motion  by  a  steam-driven  mechanism  so  as  to  produce  a 
downward  flow  of  molten  lead  in  the  sheet-iron  cylinder  a.  The  cylinder 
has  neither  top  nor  bottom,  and  being  submerged  in  the  lead,  a  circulation 
is  started,  the  lead  flowing  in  over  the  top  of  the  cylinder.  Thus  a  thor- 
ough mixing  of  the  content  of  the  kettle  is  assured.  In  a  few  minutes 


PARKES  PROCESS  505 

the  engine  is  reversed,  and  the  flow  is  made  upward  over  the  edge  of  the 
cylinder,  then  downward  to  the  bottom.  The  mixing  is  continued  about 
eleven  minutes,  after  which  time  the  stirring  apparatus  is  bodily  hoisted 
and  moved  to  one  side.  Several  kettles  can  be  thus  served  by  one  mixer. 

The  content  of  the  kettle  is  now  allowed  to  cool  two  hours  or  more. 
The  light  zinc  rises  to  the  top  and  carries  the  silver,  gold,  and  copper  with  it. 
Finally  when  the  temperature  falls  below  the  melting-point,  a  half -fused, 
mushy  crust  or  layer  forms  upon  the  lead.  The  crust  consists  of  65  per 
cent  Pb;  10  per  cent  Ag  and  Au;  3  per  cent  Cu;  and  22  to  24  per  cent  Zn. 

Fig.  274  is  an  elevation  of  the  Howard  press  by  which  the  zinc  crust  is 
removed  from  the  lead.  Another  elevation  shows  also  a  section  of  the 
cast-iron  pot  into  which  the  press  is  about  to  be  lowered.  In  principle 
the  machine  is  like  a  cheese-press.  The  apparatus  is  lowered  into  the 
lead  until  the  top  edge  of  the  cylinder  a  is  but  slightly  above  the  surface. 
The  plunger  or  follower  c  is  raised,  and  the  zinc-crust,  as  it  is  skimmed  from 
the  surface,  is  put  in  it  by  means  of  the  perforated  skimmer,  Fig.  275. 


FIG.  275.— Skimmer. 

The  press  thus  expedites  the  skimming.  While  one  man  is  skimming 
and  putting  the  skimming  into  the  press,  another  assists  by  pushing  the 
crust  to  one  place  with  a  wooden  rabble.  When  full,  the  press  is  raised 
and  the  surplus  lead  begins  to  run  out  of  the  half-inch  holes  in  the  hinged 
bottom  6.  The  plunger  c  is  brought  down,  squeezing  out  more  of  the  lead, 
and  leaves  the  remaining,  mushy,  half-fluid  mass  nearly  free  from  lead,  of 
the  composition  given  above.  The  press  is  now  run  to  one  side  over  a 
floor  paved  with  cast-iron  plates,  the  hinged  bottom  b  is  dropped  by  releas- 
ing the  catch,  and  the  zinc-crust  is  pushed  out  by  continuing  the  down- 
ward movement  of  the  plunger.  The  crust  falls  upon  the  cast-iron  floor- 
plate,  and  while  soft  is  readily  broken  with  hammers  into  lumps  the  size  of 
the  fist.  Meanwhile  the  hinged  bottom  is  closed,  and  the  press  returned 
to  the  kettle  and  is  opened  to  receive  more  skimming.  These  operations 
continue  until  the  surface  of  the  lead  is  well  skimmed  and  take  in  all 
about  twelve  hours.  The  crust  amounts  to  3000  Ib.  and  contains  90 
per  cent  of  the  silver  originally  in  the  softened  base-bullion,  resulting  in  a 
concentration  of  twenty  into  one. 

Dezincihg  Furnace.— This  first  "  zincing  "  removes  all  the  gold  and 
copper  for  which  zinc  has  a  great  affinity.  It  does  not  remove  all  the  silver, 


506 


REFINING  OF  LEAD  AND  BASE-BULLION 


and  the  operation  must  be  repeated  once  or  twice  more  before  the  silver 
content  is  diminished  to  the  fraction  of  an  ounce  per  ton  beyond  which  it 
does  not  pay  to  go.  Of  the  1.8  per  cent  zinc  needed,  first  is  added  f  of 
the  zinc  or  1.2  per  cent,  then  J,  or  0.45  per  cent,  and  finally  the  remaining 
X2-;  or  0.15  per  cent;  or  iXX),  270,  and  90  Ib.  respectively. 

The  desilverized  lead  remaining  in  the  kettle  after  the  last  skimming 
retains  0.6  to  0.7  per  cent  zinc  and  traces  of  arsenic  and  antimony,  all  of 
which  must  be  removed  before  the  lead  is  suitable  for  market.  This  is 
done  by  siphoning  or  tapping  the  metal  from  the  kettle  into  a  reverberatory 
furnace  similar  in  construction  to  the  softening  furnace.  Here  the  charge 
is^  brought  up  to  a  bright-red  heat,  the  zinc  is  volatilized  and  burned  off, 
and  litharge  forms  as  a  slag  upon  the  surface  of  the  lead.  The  operation 


FIG.  276.— Molding  Market  Lead. 

takes  six  hours,  and  is  complete  when  the  zinc  has  been  expelled,  as 
shown  by  taking  a  sample  of  lead  in  a  mold  and  observing  the  appearance 
of  the  surface  as  the  metal  solidifies.  The  furnace  is  allowed  to  cool  until 
the  litharge-slag  is  solid  and  can  be  skimmed. 

Molding. — Finally  the  lead  is  tapped  into  a  market-kettle  similar  to  the 
desilverizing  kettle.  This  is  the  reservoir  from  which  it  is  drawn  to, be  cast 
into  molds.  The  molds,  fifty  in  number,  standing  in  a  semicircle  as*  shown 
in  Fig.  276,  hold  100  Ib.  of  lead  each,  and  are  conveniently  mounted  on 
two  wheels  by  which,  when  full  and  cool,  they  are  transferred  to  the  adjoin- 
ing floor.  There  the  lead  is  tilted  out  and  the  molds  at  once  returned  to 
the  semicircle  to  be  used  again.  The  lead  is  withdrawn  from  the  kettle 
by  means  of  a  siphon.  It  descends  into  a  small  cast-iron  pot,  into  which 
is  screwed  the  2j-in.  pipe  that  delivers  it  to  the  fifty  molds,  the  pipe  being 
quickly  moved  from  mold  to  mold  as  filled,  without  interrupting  the  flow. 


RETORTING  THE  RICH  LEAD  507 

At  the  end  of  each  round  the  flow  is  interrupted  only  to  carry  back  the  end 
of  the  pipe  to  the  first  mold  of  the  series,  which  meanwhile  has  been  emptied 
and  replaced.  The  100-lb.  pigs,  or  bars,  are  the  desilverized  lead  of  com- 
merce. 

Dry  steam  may  be  blown  into  the  molten  lead  in  the  kettle  to  refine  it. 
It  is  introduced  by  means  of  a  pipe  inserted  deep  beneath  the  surface. 
The  constant  agitation  produced  by  the  steam  brings  the  metal  in  contact 
with  the  air  and  oxidizes  it  and  the  remaining  impurity.  It  is  softer  than 
ordinary  desilverized  lead,  and  is  easily  corroded  by  the  acetic  acid  used 
in  making  white  lead.  It  is  accordingly  called  "  corroding  lead." 


TREATMENT  OF  THE  RICH  LEAD 

Retorting. — Referring  to  the  diagram  (Fig.  270),  we  see  what  becomes 
of  the  crust  or  skimming  that  results  from  the  first  zincing.  The  material 
is  in  lumps,  containing  22  to  24  per  cent  zinc.  It  is  charged  with  charcoal 
breeze,  into  bottle-shaped  retorts  /,  Fig.  277,  each  holding  1200  Ib.  zinc- 
crust.  The  figure  represents  at  (a)  a  sectional  elevation  through  the  retort, 
at  (6)  a  transverse  section,  and  at  (c)  an  elevation  of  a  Faber  du  Faur 
tilting  retort-furnace.  The  retort  rests  upon  a  narrow  arch,  and  carries  a 
grate  upon  which  rests  a  coke  fire  that  fills  the  furnace  and  covers  the  retort 
/.  The  products  of  combustion  escape  by  an  outlet-port  at  the  back  to  a 
stack  of  good  draft.  The  coke  is  fed  through  a  hole  in  the  roof  of  the 
furnace,  and  is  poked  down,  and  kept  in  vigorous  combustion,  so  that  a 
yellow  heat  (1000°  C.)  is  attained.  A  condenser,  made  by -cutting  off  the 
end  of  an  old  retort  (as  shown  at  (c),  Fig.  277),  collects  the  zinc  vapors 
distilling  from  the  charge,  the  condensed  zinc  being  drawn  into  mold  sx 
through  a  1-in.  hole,  bored  through  the  bottom  edge  of  the  condenser. 
When  distillation  is  complete,  the  condenser  and  the  supporting  truck  are 
removed,  the  furnace  is  tilted  or  revolved  by  means  of  a  lever  on  the  trun- 
nions, and  the  remaining  "  rich-lead  "  is  poured  into  molds  like  those  used 
for  molding  market-lead.  The  rich  lead  still  retains  zinc,  copper,  and 
impurities,  taken  into  the  crust  at  the  first  zincing. 

Cupelling. — To  obtain  the  silver  (and  gold)  from  the  alloy,  the  English 
cupelling-furnace,  Fig.  278,  is  used.  The  principle  of  the  action  is  much 
like  that  of  cupellation  in  assaying,  except  the  litharge  here  not  only  sat- 
urates the  cupel, ^but  flows  from  it  as  fast  as  formed.  Fig.  278  shows  the 
firebox  b  where  a  long-flaming  coal  is  burned,  the  products  of  combus- 
tion passing  to  the  chimney.  The  flame  plays  over  the  hearth  called 
a  "  test,"  a  large  cupel  rammed  tight  within  a  test  ring,  shown 
mounted  on  the  carriage.  The  test  is  lowered  by  the  jack-screws,  removed 
on  the  carriage,  and  another  is  put  into  the  bottom  opening  of  the  hearth, 


508 


REFINING  OF  LEAD  AND  BASE-BULLION 

I 


t£££&£^£^^ 

(<0 

FIG.  277. — Faber  du  Faur  Retort. 


FIG.  278 — Perspective  View  of  English  Cupelling  Furnaces. 


CUPELLING 


509 


when  the  first  is  consumed.  The  test  is  hollowed  like  a  cupel,  to  hold  a 
shallow  bath  of  molten  lead  3  in.  deep.  Fig.  279  shows  two  views 
of  the  test  and  the  supporting  truck,  including  a  view  of  an  inverted 
truck  and  test.  In  the  furnace,  Fig.  278,  is  seen  the  overhead  pipe, 
branching  to  the  ash-pit,  to  supply  under-grate  blast,  and  to  an  opening 
at  the  back  of  the  furnace  where  a  tuyere  is  inserted,  by  which  a  stream  of 
air  is  brought  to  play  upon  the  surface  of  the  molten  red-hot  bath  of  rich 
lead.  The  air  oxidizes  the  lead  to  litharge.  Other  impurities  are  oxidized 
and  enter  the  litharge-slag  and  are  carried  away  with  it.  The  molten 
litharge,  as  it  forms,  escapes  by  a  shallow  groove  or  channel  in  the 
top  of  the  front  edge  of  the  cupel  or  test.  A  door  w  can  be  lifted  to 
inspect  the  operation,  or  to  cut  the  channel  as  needed.  At  the  rear  are 
provided  two  ports  of  a  size  to  permit  inserting  two  bars  of  rich  lead  that 


"?.,.?    ?    T     ?    t    5f* 

FIG.  279. — Carriage  and  Test  for  English  Cupelling  Furnace. 


are  pushed  in  as  fast  as  the  cupellation  proceeds.  The  ends  of  the  bars 
melt  and  supply  the  lead.  The  litharge  stream  is  the  size  of  a  lead-pencil, 
and  falls  into  a  small  slag-pot  beneath.  The  lead  is  fed  at  the  rate  of  1  to 
2  tons  daily  until  the  bath  has  become  rich  in  silver,  when  the  feeding  of  the 
lead  must  be  stopped.  Oxidation  then  is  continued,  cutting  the  channel 
deep  to  allow  the  remaining  litharge  to  flow  out,  and  finally  the  mirror- 
like  bath  of  silver  appears.  The  fire  must  keep  the  temperature  above  the 
melting  point  of  the  silver.  At  the  last,  a  shovelful  of  bone-ash  is  thrown 
on  the  bath  to  absorb  the  remaining  trace  of  litharge  as  it  forms  on  the 
surface.  This  is  skimmed,  and  the  silver  is  then  ready  to  be  ladled  out  or 
tapped,  commonly  into  the  cast-iron  molds,  each  holding  1000  oz.  silver. 
This  is  then  subjected  to  the  acid-parting  operation  to  be  described  later. 
The  copper-skimming,  which  is  the  first  obtained  from  the  softening 
furnace,  is  returned  to  the  blast-furnace  where  the  sulphur  of  the  charge 
combines  with  the  copper  and  removes  it  as  matte.  The  rest  of  the  skim- 
ming, mostly  lead,  containing  silver  and  gold,  is  reduced  to  base-bullion. 


510  REFINING  OF  LEAD  AND  BASE-BULLION 

The  third  skimming  of  the  softening-furnace,  if  any,  is  returned  to  the 
blast-furnace,  since  it  contains  but  little  antimony. 

The  second  softening  skimming  or  antimony-skim,  containing  15  to 
25  per  cent  antimony,  goes  to  a  small  reverberatory  furnace,  8  by  12  ft. 
hearth  dimensions  and  10  in.  deep,  built  like  a  softening-furnace  and 
called  a  precipitating-furnace.  Here  it  is  melted,  with  a  reducing  flame 
into  a  slag.  Charcoal  is  added,  and  stirred,  to  reduce  or  precipitate  part 
of  the  lead  of  the  slag.  The  lead,  falling  to  the  bottom  of  the  bath,  carries 
down  the  silver  of  the  slag.  When  the  reaction  is  complete,  the  super- 
natant slag  is  tapped  into  slag-pots,  and  the  lead  is  tapped  into  a  kettle 
at  a  lower  level  and  molded  into  bars.  This  precipitated  lead-bullion  is 
returned  to  the  softening-furnace  to  be  softened  and  desilverized.  The 
antimonial  slag,  containing  about  6  oz.  silver  per  ton,  when  accumu- 
lated, is  smelted  in  a  small  blast-furnace  to  reduce  it  to  antimonial  lead  of 
20  per  cent  Sb,  which  is  sold  to  the  type  founders.  The  slag  is  rejected. 

THE  PATTINSON  PROCESS 

When  a  kettle  containing  molten  lead  is  allowed  to  cool  slowly  as  it 
approaches  solidification,  crystals  of  lead  low  in  silver  separate.  The 
metal  that  remains  liquid  contains  the  larger  part  of  the  precious  metal. 
The  crystals  are  removed  with  a  perforated  ladle,  melted  in  another  kettle, 
and  allowed  to  cool.  Once  more  crystals  separate  that  are  low  in  silver, 
the  mother  liquor  becoming  high  in  silver.  If  the  liquid  portion  first  men- 
tioned be  transferred  to  a  kettle  and  likewise  heated  and  then  allowed  to 
cool,  the  same  segregation  of  the  silver  into  the  liquid  part  continues. 
We  can  accordingly  arrange  a  series  of  kettles  containing,  at  one  end  low- 
grade  lead,  and  at  the  other  high-grade,  all  from  one  product.  A  series 
of  this  kind,  as  illustrated  by  practice  at  Eureka,  Nev.,  gave  the  assays 
quoted  in  the  table  below. 

Another  crystallization  would  reduce  the  silver  of  the  market  lead  to 
half  the  value  given.  The  rich  lead  could  be  directly  cupelled  in  an 
English  cupelling-furnace,  or  better,  treated  by  the  Parkes  process  to  get 
rich  silver-zinc  crust  for  retorting  and  cupelling.  The  process  has  been 
modified  recently  by  Tredennick,  who  raises  each  kettle  by  hydraulic 
power  above  the  adjoining  one  so  that  the  mother  liquor  drains  fr$>m  one 
kettle  to  the  next  through  a  strainer,  the  lead  being  cooled  near  the  solidi- 
fication temperature  by  introducing  steam  upon  the  surface.  The  cost  of 
operating  has  been  greatly  decreased  in  this  way.  The  chief  advantage  of 
the  Pattison  process  over  the  Parkes  is  that  it  gives  a  product  free  from 
bismuth.  In  the  Parkes  process  the  bismuth  follows  the  lead.  Bismuth 
is  injurious  in  lead  that  is  to  be  corroded  to  make  white  lead,  and  it  may  be 
necessary  to  employ  the  Pattinson  process  for  making  a  corroding  lead  from 
bismuth-bearing  ores. 


BETTS  PROCESS  511 

Market  Lead 
Kettle  Ounces  Ag  per  Ton. 

No.     1 1.25 

No.    2 2.5 

No.    2 5.0 

No.    4 9.0 

No.    5 18.0 

No.    6 30.0 

No.    7 50. 0 

No.    8 75.0 

No.    9 100. 0 

No.  10 150 . 0 

No.  11* V.  -'I?. 450.0 

*  Rich  Lead. 

COST  OF  REFINING  BASE-BULLION 

The  actual  cost  x)f  refining  base-bullion  is  as  follows: 

Prime  or  flat-cost,  of  softening  and  refining $5.00  to    $6.00 

General  expense : 3.00  to      3.00 

Loss  in  metals  and  incidental  expenses 1 . 70  to      3 . 00 


Total $9.70  to  $12.00 

SELLING  PRICE  OF  BASE-BULLION 

Base-bullion. — A  small  works  may  sell  this  product  to  an  Eastern 
refinery  as  follows : 

Gold  is  paid  for  at  $20.40  per  ounce.  Silver  at  99  per  cent  of  New  York 
quotation,  sixty  days  after  date  of  sampling  at  the  refinery.  Lead  is  paid 
for  at  99  per  cent  of  the  New  York  quotation,  thirty  days  after  same  date 
of  sampling. 

Refining  cost,  $12  per  ton.  An  advance  will  be  given  up  to  90  per  cent 
of  the  net  value  with  a  charge  of  6  per  cent  for  said  advance. 


THE  BETTS  PROCESS  FOR  THE  ELECTROLYTIC  REFINING  OF  LEAD 

The  principle  of  this  process  depends  upon  the  solubility  of  lead  in  an 
acid  solution  of  lead  fluosilicate,  which  is  used  as  an  electrolyte.  The 
solution  is  formed  by  diluting  hydrofluoric  acid  containing  35  per  cent  HF 
with  an  equal  volume  of  water  and  saturating  with  powdered  quartz 
according  to  the  reaction : 

(9)  SiO2+6HF  =  H2SiF6+2H2O. 

In  the  hydrofluosilic  acid  lead  is  dissolved.  The  solution  contains 
7  to  10  per  cent  lead,  and  8  to  12  per  cent  of  fluosilicic  acid  (H2S;  F6),  the 
free  acid  varying  from  3  to  5  per  cent  lead.  To  obtain  a  solid  deposit  on 
the  cathode  glue  is  added  to  the  extent  of  0.1  per  cent. 


512  REFINING  OF  LEAD  AND  BASE-BULLION 

The  anodes  are  plates  of  the  base-bullion  to  be  refined,  cast  1J  in.  thick, 
resembling  ordinary  copper  anodes. 

The  cathode-sheets  that  receive  the  deposited  lead  are  "  stripping 
plates/'  obtained  as  in  the  case  with  copper  cathodes.  They  are  made 
by  depositing  lead  upon  steel  cathode-plates,  prepared  for  use  by  cleaning, 
coating  with  copper,  lightly  lead-plating  them  in  the  tanks,  and  greasing 
with  paraffin.  On  them  is  deposited  the  lead,  and  when  the  coating  is  of 
the  desired  thickness  the  steel  cathodes  are  removed  from  the  bath,  and 
the  lead  coating  or  sheets  are  stripped  off  for  use  as  cathode.  Another 
method  consists  in  casting  the  cathodes  in  the  form  of  thin  sheets. 

The  anodes  and  cathodes  are  placed  If  in.  apart  in  the  tank.  As 
in  copper  refining,  the  anodes  are  in  multiple,  and  the  tanks  in  series. 
The  current  enters  the  anodes,  passes  through  the  electrolyte  to  the 
cathodes,  dissolves  the  lead  from  the  anodes  and  deposits  it  upon  the 
cathodes. 

The  fall  of  potential  between  anode  and  cathode  is  0.45  volt,  and  the 
current  strength  is  15  to  18  amperes  per  square  foot.  One  ampere  deposits 
3J  oz.  lead  per  twenty-four  hours  following  the  ratio  of  the  atomic  weights 
of  copper  and  lead  which  is  63.6  to  207. 

In  the  process  the  impurities  remain  as  an  adherent  coating  on  the 
anode,  and  consist  of  the  copper,  bismuth,  arsenic,  gold,  and  silver.  The 
zinc,  iron,  cobalt,  and  nickel  dissolve  in  the  electrolyte. 

The  cathode  (containing  the  starting  sheet)  is  melted  and  cast  into  bars; 
the  anode  mud  recovered  from  the  anode  and  collected  at  the  bottom  of 
the  tank  is  refined  to  recover  the  precious  metals.  When  about  three- 
fourths  corroded  the  anode  is  removed. 

As  compared  with  ordinary  refined  lead,  electro] ytically  refined  lead  is 
pure,  being  practically  free  from  bismuth,  even  when  much  is  present  in 
the  base-bullion,  and  it  must  be  remembered  that  bismuth  is  harmful  to 
"  corroding  lead." 

The  residue  or  anode-slime,  averaging  8000  oz.  or  more  of  silver  and  gold 
per  ton,  is  treated  by  boiling  it  with  sulphuric  acid,  using  a  steam  pipe 
inserted  in  the  solution  to  boil  and  agitate  it  with  free  access  of  air.  The 
washed  residue  is  melted  in  a  small  basic-lined  reverberatory  furnace,  the 
copper  is  removed  by  using  niter  as  a  flux,  and  the  antimony  by  the  addi- 
tion of  soda.  The  dore  bars  finally  obtained  are  parted  in  the  usual  way 
with  sulphuric  acid. 

The  process  is  used  at  East  Chicago  and  Trail,  B.  C.  At  Omaha 
it  is  confined  to  the  refining  of  base  bullion  containing  1.5  per  cent 
bismuth. 


SILVER-LEAD  SMELTING  WORKS 


513 


SMELTERY  AND  REFINERY  FOR  SILVER-LEAD  ORES 

We  show  in  Fig.  280  a  general  view  of  a  company  plant  for  the  treat- 
ment of  their  own  ores  and  base-bullion. 

Referring  to  Fig.  280  a  double  track  at  the  right  (in  this  case  the  West 


side)  brings  in  coke  and  ore,  the  first  to  the  coke  storage  and  the  second 
to  the  ore  storage  bins  at  the  extreme  left.  Here  the  coarse  ore  is  taken 
out  to  be  sent  to  the  blast-furnace  building,  and  the  fines  go  to  the  Wedge 


514  REFINING  OF  LEAD  AND  BASE-BULLION 

roasters  and  the  Dwight-Lloyd  sinter  plant.  The  base-bullion  from  the 
blast-furnace  adjoining  goes  to  the  lead  refinery,  giving  two  products,  the 
market  lead  to  the  lead  storage  warehouse  and  dore  bars  to  be  parted  at 
the  gold  and  silver  refinery.  The  blast-furnace  flue,  parallel  to  the  blast- 
furnace building,  turns  at  right  angles,  forming  the  "  trail  "  to  the  fan- 
house  where  the  fume  is  driven  through  the  bag-house  and  finally  to  the 
15-foot  by  200-foot  Custodis  stack.  Powdered  coal  is  used  in  the  refinery 
and  the  Dwight-Lloyd  sinter  plant. 


PART  VII 
ZINC 


CHAPTER  XL 

ZINC  AND  ITS  ORES 

PROPERTIES  OF  ZINC 

Zinc  is  a  bluish-white  metal  of  specific  gravity  6.9  to  7.2,  according  to 
the  way  it  has  been  cast  and  cooled.  The  rolled  metal  has  a  specific  gravity 
of  about  7.25.  Zinc  melts  at  419°  C.  and  boils  at  950°  C.  with  a  charac- 
teristic brilliant  bluish-green  flame.  The  commercial  metal  becomes 
malleable  and  ductile  if  heated  to  100°  to  150°  C.  and  when  cooled  from 
this  can  be  rolled.  At  205°  C.  it  again  becomes  brittle,  and  may  be  pul- 
verized in  an  iron  mortar.  Zinc  has  a  tensile  strength  of  but  18,700  to 
22,200  Ib.  per  sq.  in.  when  in  sheets  or  wire.  When  it  passes  from  the 
cold  solid  to  the  molten  condition  it  increases  in  volume  and  on  again 
cooling  contracts  but  slightly.  Carbon  dioxide  readily  oxidizes  zinc 
vapor  with  the  production  of  carbon  monoxide  and  zinc  oxide. 

To  the  metallurgist  ks  value  in  the  recovery  of  the  precious  metals  in 
the  Parkes  process  and  its  usefulness  in  cyanidation  largely  appeals. 

ZINC  ORES 

The  principal  ores  of  zinc  are  blende  and  calamine.  In  New  Jersey 
occur  deposits  of  franklinite,  ZnOFe2Os.  The  zinc  minerals  seldom  occur 
pure;  besides  the  earthy  gangue,  and  sulphides  of  iron,  lead  and  copper, 
blende  as  marmatite  or  black-jack  contains  iron  so  combined  chemically 
that  it  cannot  be  separated  by  ore-dressing  methods. 

Blende,  or  sphalerite  (ZnS),  when  of  a  yellow  color  as  in  the  ore  of  the 
Joplin  district  in  Missouri,  is  called  rosin-blende.  When  dark  in  color, 
due  to  chemically  contained  iron  as  in  the  ore  of  the  Rocky  Mountain 
States,  it  is  called  black-jack.  It  is  from  blende  concentrate  that  most  of 
the  spelter  of  commerce  is  extracted.  It  needs  roasting  before  it  can  be 
retorted  or  smelted  for  extracting  the  zinc. 

Calamine  is  a  term  applied  commercially  both  to  the  carbonate  (smith- 
sonite)  and  to  the  hydrous  silicate  of  zinc.  It  is  an  oxidized  or  sulphur- 
free  ore  that  needs  no  preliminary  roasting  before  smelting.  On  being 
heated  in  the  retort,  the  CO2  of  the  carbonate  is  expelled,  leaving  zinc 
oxide. 

Willemite,  the  anhydrous  silicate  occurring  with  franklinite  in  New 

517 


518  ZINC  AND  ITS  ORES 

Jersey,  mixed  with  coal,  is  decomposed  at  the  high  temperature   of   the 
retort,  yielding  zinc. 

It  is  generally  found  advantageous  to  calcine  calamine  for  the  purpose 
of  driving  off  CO2  and  water,  which  are  undesirable  in  retorting  because  of 
their  oxidizing  action  on  zinc-vapor.  However,  the  preliminary  calcina- 
tion is  often  omitted,  but  when  performed,  it  is  done  in  kilns  much  like 
those  in  which  lime  is  burned. 


CHAPTER   XLI 
ROASTING  ZINC  ORES 

REDUCTION  OF  ORES  OF  ZINC 

In  outline,  the  metallurgy  of  zinc  consists  in  grinding  the  ore  (gener- 
ally blende)  and  roasting  to  convert  into  zinc  oxide,  then  charging  the 
roasted  ore,  ultimately  mixed  with  fine  coal,  into  horizontal,  cylindrical, 
clay  retorts,  heated  to  a  white  heat,  where  the  zinc,  reduced  by  the  coal, 
volatilizes,  and  the  vapor,  entering  the  cool,  tapering,  clay  extension  of  the 
retort  (called  the  condenser),  condenses  there.  As  it  accumulates  it  is 
tapped  into  a  ladle  from  time  to  time,  skimmed,  and  cast  in  molds. 
When  distillation  is  complete  the  condenser  is  removed  and  the  content  of 
the  retort  taken  out  and  generally  thrown  away.  The  cycle  of  opera- 
tions takes  twenty-four  hours. 

ROASTING  BLENDE 

The  aim  is  to  dead-roast  the  ore,  generally  to  1  per  cent  sulphur  or  less. 
For  every  1  per  cent  sulphur  remaining  in  the  roasted  ore,  2  per  cent  zinc 
is  held  back  in  the  retort-residue  after  distillation. 

The  ore  is  ground  to  about  6-mesh  size,  then  slowly  and  carefully 
roasted  with  frequent  stirring,  finishing  the  roast  at  a  high  temperature  to 
decompose  the  zinc  sulphate  formed  at  the  lower  temperature.  The  ore  is 
generally  in  the  form  of  concentrate,  still  containing  a  little  gangue,  galena, 
and  pyrite.  To  remove  the  final  1  per  cent  of  sulphur  would  require  a  long 
time  and  would  not  be  commercially  profitable.  The  ore  is  accordingly 
considered  to  be  finished  when  it  contains  no  more  than  that  amount  of 
sulphur. 

CHEMISTRY  OF  ROASTING  ZINC   ORES 
We  have,  in  the  roasting  of  blende,  the  following  reactions: 

(1)  ZnS  +  40   =   ZnO  +    SO3 

43,000  86,400      71,000=  +114,400  cal. 

(2)  2ZnS    '+  7O  =  ZnO  +    ZnSO4   +   S02 

2X43,000  86,400       230,000      7 1,000  =+30 1,400  cal. 

519 


520  ROASTING  ZINC  ORES 

Thus  in  an  oxidizing  flame,  blende  is  roasted  to  oxide  and  sulphate,  both 
reactions  being  exothermic.  As  indicated  in  the  reactions  given  in  the 
chapter  on  Roasting,  pyrite  or  chalcopyrite  assists  in  the  reactions.  At  a 
cherry-red  heat  the  zinc  sulphate  is  decomposed  into  basic  sulphate 
(3ZnO,  ZnSO4)  thus: 

(3)  4ZnSO4  =  3ZnO,  ZnSO4+3SO3. 

The  basic  sulphate,  exposed  to  a  bright-red  heat  for  a  time,  reacts  thus: 

(4)  3ZnO,  ZnSO4  =  4ZnO +SO3. 

Finally  zinc  oxide  is  obtained  and  the  sulphuric  anhydride  is  eliminated. 

When  limestone  or  calcite  is  present  it  is  converted  in  large  part  to 
sulphate.  Galena  also  roasts  to  a  sulphate,  and  tends  to  envelop  particles 
of  blende,  and  to  prevent  their  roasting.  Much  of  the  blende  from  Lead- 
ville,  Colo.,  and  other  Western  States,  contains  silver,  and  it  consequently 
often  pays  to  treat  the  retort-residues  after  the  zinc  has  been  removed. 

There  is  a  loss  of  silver  in  roasting  that  may  be  given  at  10  to  15 
per  cent  and  also  a  loss  of  zinc  as  dust  and  volatilization  at  the  final 
high  temperature  that  may  be  reckoned  at  2  per  cent  or  more. 

In  roasting  zinc  sulphides,  and  especially  flotation  concentrates,  better 
results  in  retorting  are  attained  by  finishing  the  roast  without  an  excess  of 
air.  Thus  roasted,  the  product  sinters  and  becomes  more  porous,  and 
the  sulphur  left  in  the  ore  is  Jess  harmful. 

ROASTING  FURNACES 

The  roasting  of  blende  has  been  performed  in  hand-rabbled  reverbera- 
tory  furnaces  as  well  as  in  a  great  variety  of  mechanical  furnaces.  These 
are  described  in  the  chapter  on  Roasting.  The  latter  are  gradually  sup- 
planting the  former  because  of  the  saving  of  labor.  It  should  be  noted, 
however,  that  the  wear  is  great  on  mechanical  furnaces  that  have  ironwork 
exposed  to  the  heat  because  of  the  high  final  heat  needed  in  blende-roasting 
and  consequently  the  types  of  furnace  have  been  preferred  where  the  rabble 
is  exposed  but  a  short  time  to  the  action  of  the  fire,  and  where  iron  parts  are 
not  exposed  or  can  be  water-cooled. 

Thus,  the  Brown  horseshoe  furnace,  where  the  rabble  is  drawn  through  a 
circular  hearth,  then  allowed  to  cool,  or  the  Wethey  furnace,  where  the 
rabble  is  exposed  to  the  fire  but  half  of  the  time  and  the  moving  iron  parts 
are  outside  the  furnace,  have  been  successfully  used  in  blende-roasting. 
Of  the  recent  types,  the  Hegeler  furnace  has  proved  most  successful  for 
the  above  reasons.  It  is  a  multiple-hearth  furnace  closed  by  swinging 
sheet-iron  doors  at  the  ends,  and  stirred  by  rabbles  drawn  quickly  through 
the  furnace  by  means  of  rake  rods,  so  that  the  parts  are  outside  the  furnace 


THE  WEDGE  ROASTING  FURNACE  521 

most  of  the  time  and  no  iron  parts,  except  the  end  swinging  doors,  are 
affected  by  the  fire.  The  hearths  being  superimposed  make  a  compact 
furnace,  and  the  radiation  is  greatly  lessened,  so  that  there  is  economy  of 
fuel. 

THE  WEDGE  MECHANICAL  BLENDE-ROASTING  FURNACE 

This,  one  of  the  most  successful  of  the  blende  roasters  where  it  is  desired 
to  save  the  sulphur  fumes  for  sulphurous  acid,  is  shown  in  Fig.  281. 

It  is  commonly  used  for  an  oxidizing  roast,  the  gases  carrying 
about  2^  per  cent  sulphur,  but  for  use  in  making  sulphuric  acid  the 
gas  should  carry  6  to  7  per  cent  of  sulphur.  This  is  thus  accomplished: 
The  blende  carrying  about  25  per  cent  of  sulphur  is  self-roasting,  the  sul- 
phur escaping  at  the  desired  strength.  Below  the  fifth  hearth,  however, 
when  no  more  than  8  per  cent  sulphur  remains,  it  needs  a  fire  to  roast  it. 
The  floor  of  the  fifth  hearth  is  thin  and  between  it  and  the  roof  of  the  sixth 
hearth  is  a  muffle  space  where  the  flame  from  the  firebox  at  the  left  enters 
and  is  down-drafted.  That  is,  the  flame  passes  through  drop-holes,  then 
successively  through  the  sixth  and  seventh  hearths  to  a  side-flue  leading 
to  the  chimney,  and  in  its  progress  roasting  the  ore  from  8  per  cent  down  to 
1  or  2  per  cent,  the  degree  needed  for  properly  roasted  ore.  In  the  thick- 
ness of  the  side  walls  is  a  drop-hole  or  passage  downward  debouching  into 
No.  6  hearth.  Ore  pushed  into  this  drop-hole  fills  it  with  a  talus  at  the 
outlet  which  is  swept  away  by  the  rabble  of  No.  6  hearth.  Thus  the  ore 
passes  down  by  a  sealed  opening  while  the  fire  gases  cannot  escape  upward, 
To  mingle  with  the  strong  sulphur  fumes  produced  above. 

THE  HEGELER  FURNACE 

Fig.  282  gives  a  transverse  section,  a  longitudinal  elevation  and  section 
and  a  plan  of  the  furnace,  of  75  ft.  effective  length,  and  built  double. 
Referring  to  one  side  there  are  seven  roasting-hearths  each  marked  B, 
three  fire-hearths  A,  and  a  hearth  C,  for  preheating  the  air  for  the  roasting 
hearth.  Thus  the  three  lower  roasting-hearths  B,  are  heated  by  the  fire- 
hearths,  constituting  muffles,  so  that  the  fire-gases  do  not  mingle  with  the 
roaster  gas.  In  operation,  the  furnace  being  at  full  heat,  the  ore,  to  the 
amount  of  one  charge  is  dropped  from  the  hopper  D  upon  the  upper  or 
No.  1  hearth.  By  the  action  of  a  rake  of  the  width  of  the  hearth,  intro- 
duced through  the  end  door  at  that  end,  it  is  pushed  along  and  spread  out 
upon  the  just-emptied  hearth.  The  next  time  the  rake  is  passed  through 
the  ore  is  worked  to  the  opposite  end  of  No.  1  hearth  and  through  the  open- 
ing E,  to  the  second  hearth  B.  Another  rake  propels  it  through  B,  and 
so  progressively  through  the  remaining  hearths  until,  at  the  lowest  one,  it 
discharges  through  a  side-opening  into  a  waiting  ore  car.  The  path  of 


522 


ROASTING  ZINC  ORES 


SECTION 


ELEVATION 
Sliding  Drop  Hole 


PrivingPu 


SECTION 


PUN- 


.  281.— Wedge  Roaster  for  Zinc  Ores. 


THE  HEGELER  ROASTING  FURNACE 


523 


the  fire-gases  is  in  an  opposite  direction,  the  producer  or  natural  gas  being 
admitted  to  the  lowest  hearth  A,  at  one  end,  and  traveling  to  the  other, 
being  partly  burned  by  the  admission  of  air  at  side-ports.  The  gas  then 
by  a  by-pass  at  the  side  goes  to  the  hearth  A,  there  to  be  further  burned,  and 
the  combustion  is  finished  on  the  third  or  upper  fire-hearth  A .  At  the  end 
of  this  the  gases  pass  into  a  flue  of  their  own,  and  thence  to  the  chimney. 
Air  for  roasting  is  admitted,  partly  in  the  lower  hearth  B,  partly  in  the  two 
hearths  above  it,  and  taking  a  course  opposite  to  that  of  the  ore  in  its 
descent,  leaves  the  top  of  the  furnace  at  the  opening  F,  into  the  flue  G. 
It  will  be  observed  that  the  furnace  is  built  double,  so  that  when 


I T 


T      I      T 


PART  SIDE  ELEVATION 


FIG.  283. — Hegeler  Roasting  Furnace. 

ore  is  admitted  at  D,  it  is  also  admitted  at  X  on  the  opposite  end,  and  the 
same  rake  works  both  hearths  in  succession.  The  rakes  or  other  moving 
parts  do  not  remain  constantly  in  the  furnace,  but  a  rod  is  passed  through 
from  one  end,  hooks  on  to  the  rake  and  pulls  it  back  with  it. 

The  furnace  is  used  in  this  country  where  it  is  necessary  to  convert  the 
sulphur  gases  into  sulphuric  acid,  but  on  account  of  the  high  labor  and 
maintenance  costs,  is  rarely  used  where  acid  is  not  made.  The  capacity 
of  the  furnace  is  about  45  tons  per  day,  with  a  coal  consumption  of  25  to 
35  per  cent  of  coal  dependent  on  conditions.  The  furnace  is  also  sometimes 
built  with  regenerative  chambers  for  the  fire  gas,  whereby  the  fuel  consump- 
tion .is  reduced  where  the  high  price  of  fuel  makes  this  imperative. 


524 


ROASTING  ZINC  ORES 


THE  MERTON  ROASTING  FURNACE  525 


VARIOUS  FURNACES 

Other  furnaces  used  for  the  roasting  of  blende  where  acid  is  not  made 
from  the  waste  gases,  are  the  Zellweiger  and  the  Ropp,  both  of  them 
straight  line  single  hearth  furnaces  of  the  Brown-O'Hara'type.  In  Euro- 
pean practice  the  Merton  and  the  Ridge  furnaces  are  somewhat  generally 
adapted  for  blende  roasting,  although  on  account  of  cheaper  labor,  hand 
roasters  are  somewhat  common  there. 

THE  MERTON  FURNACE 

This  is  a  muffled  furnace,  having  six  muffles,  where  the  ore  is  roasted,  and 
between  the  floor  of  one  and  the  roof  of  the  next  a  space  transversed  by  the 
fire-gases.  Referring  to  the  longitudinal  section,  Fig.  384,  the  ore,  deliv- 
ered by  a  feeder  (not  shown)  at  the  left  end,  falls  upon  the  No.  1  hearth,  and 
by  means  of  eight  rabbles  is  handed  along  to  the  other  end  where  it  drops 
upon  No.  6  hearth  and  escapes  by  a  drop-hole  into  a  car  standing  upon  a 
track  below.  It  will  be  seen  that  No.  3  hearth  is  muffled  neither  above  nor 
below.  The  fumes  from  the  three  top-hearths  go  to  the  sulphuric  plant, 
the  fire-gases  entering  the  muffled  spaces  between  hearths  4  and  5  and 
between  5  and  6  are  down<lrafted  to  the  chimney,  while  the  muffle  gases 
between  1  and  2  go  direct  to  the  stack.  In  this  way  the  ore  is  soon  heated 
to  ignition,  and  later,  on  5  and  6  is  strongly  heated  to  give  a  low  sulphur 
product.  It  will  be  noticed  that  the  muffle  roofs  are  so  low  that  the  gases 
traveling  over  the  hearth  continually  sweep  along  the  escaping  862  gas. 
On  hearth  No.  6  there  is  no  muffle  below  so  that  the  ore  becomes  cooler 
and  in  so  doing  heats  the  entering  air. 

THE  RIDGE  FURNACE 

Fig.  285  shows  three  sections — one  a  central  longitudinal,  one  trans- 
verse, through  B,B,  and  one  as  a  plant  at  B,B,  cutting  through  hearth  No. 
3.  The  upper,  the  drying  and  preheating  hearth,  is  open  to  the  air,  and 
the  roasting  of  the  ore  is  done  on  hearths  Nos.  1,  2,  and  3,  passing  thence 
to  the  cooling  hearth.  Below  hearth  No.  3  are  flues  for  fire-gases  and  for 
fresh  air,  which  pass  away  by  a  separate  flue.  In  this  way  hearth  No.  3  is 
maintained  at  a  high  temperature,  being  a  muffle  hearth.  After  the  fur- 
nace has  been  heated  to  a  high  temperature  at  starting,  the  blende  is  self- 
burning,  air  for  the  purpose  being  admitted  by  the  proper  side-ports  as 
shown  on  the  plan.  The  gas  evolved  from  the  burning,  and  containing  6J 
to  8|  per  cent  SO2  (free  from  fire  gases)  passes  by  the  round  gas  flues,  and 
through  a  suction  fan,  to  the  sulphuric-acid  plant.  The  direction  of  flow 
of  this  gas  is  contrary  to  the  movement  of  the  ore.  There  are  four  ver- 
tical hollow  shafts  carrying  rabble  arms  furnished  with  rabbles  or  blades 


526 


ROASTING  ZINC  ORES 


SULPHURIC  ACID  MAKING  527 

set  at  an  angle.  The  shafts  being  in  motion,  ore  from  the  feed  is  swept 
along  by  the  rabbles  the  length  of  the  drying  and  preheating  hearth.  It 
falls  through  the  open  drop-hole  at  the  firebox  end  of  the  furnace  to  roasting 
hearth  No.  1,  thence  along  the  hearth  to  the  drop-hole  delivering  to  hearth 
No.  2  and  so  on  to  hearth  No.  3  and  to  the  cooling  hearth.  The  discharge 
is  at  the  side  of  the  hearth  and  outside  the  brick  wall  which  protects  the 
gearing  from  dust  and  heat. 

SULPHURIC  ACID 

This  is  a  by-product  of  zinc  roasting.  The  sulphur  dioxide  fumes  arising 
from  blende  roasting  are  increasingly  used  for  the  manufacture  of  sulphuric 
acid.  This  is  done  not  only  because  of  the  profit  arising  from  this  manu- 
facture, but  because  escape  of  these  fumes  into  the  atmosphere  results  in 
damage  to  health  and  to  vegetation,  so  that  such  disposal  of  fumes  has 
been  restricted  by  legislation.  Most  blende  is  free  from  arsenic,  and  the 
fume  arising  from  its  roasting  makes  a  superior  acid  as  compared 
with  that  made  from  pyrites.  Due  to  this  freedom  from  arsenic  the  con- 
tact process  for  making  acid  is  much  in  use  in  the  United  States,  though 
the  older  chamber  process  is  employed,  especially  for  the  Western  blendes 
carrying  iron  and  lead.  One  may  note  hi  connection  that  it  has  been 
found  to  be  an  advantage  to  roast  the  blende  in  one  establishment  near  the 
zinc  mines  and  the  sulphuric  acid  market,  and  to  smelt  the  roasted  product 
at  another  works  where  fuel,  clay,  and  skilled  labor  are  best  found.  To 
roast  blende  to  the  best  advantage  the  operation  should  be  carried  on  in  a 
multiple-hearth  muffle-furnace  where  one  can  be  sure  of  a  regular  supply  of 
sulphur  dioxide,  free  from  the  combustion  gases,  and  of  suitable  grade. 
The  objection  to  hand-roasting  is  production  of  an  irregular  gas  due  to 
uneven  firing  and  stirring,  whereas  in  the  mechanical  furnace  such  condi- 
tions do  not  exist.  Again,  in  a  muffle  furnace  the  fuel  gases  are  kept 
separate  from  the  sulphur  fume  arising  from  the  roasting  ore,  and  these 
may  be  maintained  at  the  grade  of  from  6J  to  8J  per  cent  sulphur 
as  best  suited  for  acid  making.  In  1920  the  average  price  of  60°  Baume 
acid  was  $14  to  $18  per  ton. 


CHAPTER  XLII 
SMELTING  ZINC  ORES 

THE  SMELTING  OR  DISTILLATION  OF  ROASTED  ZINC  ORES 

The  recovery  of  zinc  from  the  ore  consists  in  distillation  of  the  roasted 
ore  in  refractory  clay  retorts  after  intimately  mixing  it  with  40  to  60  per 
cent  of  its  weight  of  fine  coal.  The  whole  is  brought  to  a  white  heat,  which 
is  maintained  during  an  entire  day. 

Reactions  that  Occur  in  Retorting  Roasted  Zinc  Ore. 

(5)  ZnO  +  C   =  Zn  +  CO. 

86,000  29,000   =  +  57,000 

(6)  2ZnO  +  C .  =  2ZnCO2. 
172,000  97,000  =+57,000 

(7)  C02  +  C   =  2CO. 

97,000  58,000  =-39,000 

(8)  ZnO  +  CO     =  Zn  +  CO2. 

86,000    29,000  97,000   =-18,000 

Before  ore  and  coal  are  charged  into  the  hot  retort  the  mixture  is 
moistened  for  convenience  in  charging,  the  water  being  promptly  driven 
off  by  the  heat.  The  light  hydrocarbons  of  the  coal  come  away  next; 
then  the  iron  oxide  is  reduced  to  protoxide  and  part  of  it  to  a  porous  iron 
or  iron  sponge.  The  final  reaction  (6)  is  the  reduction  of  the  zinc  oxide  of 
the  ore  by  the  carbon  to  metallic  zinc.  The  reaction  commences  at 
1060°  C.,  but  practically  a  temperature  of  1300°  C.  is  reached. 
*  It  is  to  be  noted  that  the  reduction  point  of  zinc  oxide  is  considerably 
above  the  boiling-point  of  zinc  so  that  metal  so  reduced  in  the  retort  is 
immediatly  carried  off  with  the  other  products  of  reduction  into  the  cooler 
condenser.  Should  the  temperature  of  this  condenser  rise  higher  than  the 
boiling-point  (circa  975°  C.)  the  zinc  will  of  course  fail  to  condense,  and 
should  it  fall  below  the  melting-point  (circa  418°  C.)  the  metal  will  be  con- 
densed as  powder,  known  generally  as  blue  powder.  This  later  character- 
istic is  taken  advantage  of  in  the  manufacture  of  zinc  powder  or  blue 

528 


ZINC  SMELTING  FURNACE 


529 


FIG.  286. — Zinc-smelting  Furnace. 


powder — now  so  much  used  in  the  reduction  of  gold  from  cyanide  solutions 

by  using  iron  condensers  maintained  at  a  sufficiently  low  temperature. 

The  condensers  employed  for  spelter  are  truncated  hollow  cones  of  fireclay, 
which  fit  just  inside  the  mouth  of  the  retort.     These  are  generally  made 


FIG.  287. — Old  Furnace  of  Belgium  Type. 


530 


SMELTING  ZINC  ORES 


about  18-24  in.  long  and  taper  to  about  4  in.  outside  diameter  at  the 
smaller  end,  with  walls  about  J  in.  thick. 

In  the  United  States  the  Belgian  style  retort  is  chiefly  used.  These 
retorts  are  plain  cylindrical  vessels,  closed  at  one  end,  and  are  made 
48  to  54  in.  long,  and  11  in.  outside  diameter.  The  thickness  of 
the  side  walls  is  1  to  1J  in.  and  of  the  end  about  1J  in.  The  older 
style  of  zinc  furnace  designed  for  the  direct  use  of  coal  as  fuel, 
and  known  generally  in  this  country  as  the  Belgian  furnace,  is  shown  in 


FIG.  288. — Section  of  Zinc-smelting  Furnace  (gas-fired). 

cross-section  in  elevation  in  Fig.  286.  Fig.  287  also  shows  two  photo- 
graphic views  of  this  same  type  of  furnace.  On  account  of  their  high 
labor  cost  in  operating  these  furnaces  are  rarely  used  to-day.  Instead  the 
larger  and  lower  furnace  shown  in  cross-section  Fig.  288  and  in  front 
elevation  Fig.  289  is  more  generally  employed  both  for  the  use  of  natural 
gas  and  for  producer  gas  as  fuel.  The  furnaces  are  built  with  any  number 
of  retorts,  generally  288  to  336  to  a  side  where  four  retorts  high.  It  is 
common  where  producer  gas  is  used  as  fuel  to  build  them  five  and  even  six 
rows  high,  in  which  case  the  furnaces  are  made  400  retorts  and  432  retorts 


ZINC-SMELTING  FURNACE 


531 


to  a  side.  Where  producer  gas  is  employed  fuel  producers  are  located 
on  the  end  of  the  furnace,  and  all  the  gas  is  allowed  to  enter  at 
that  end.  Air  is  supplied  at  intervals,  by  blowers  or  fans,  through  the 
pipes  A  shown  on  top  of  the  furnace  and  distributed  through  the  smaller 
pipes  B  to  the  front  at  each  section.  The  gases  pass  out  through  stacks 
located  at  the  end  opposite  to  the  producers. 

Where  natural  gas  is  used  for  fuel,  no  producer  is  of  course  employed, 


r*-i8 


00 


oo 


00 


oo 


oo 


oo 

V  J  \^J 


oooc 


oo 


oo 


oo 


oo 


o 


rr 


^x£pL    j^gXi  pa 

11  '»!  ^  -m«T  i    ' »j  '5' 


FIG.  289. — Front  View  of  Smelting  Furnace  (gas-fired). 

and  that  end  of  the  furnace  is  closed  up.  The  gas  is  admitted  at  intervals 
along  the  furnace  with  the  air,  and  the  products  of  combustion  pass  out  the 
stack  on  one  end  of  the  furnace. 

Both  these  styles  of  furnace  are  extremely  wasteful  of  fuel,  as  the 
products  of  combustion  leave  the  furnace  at  the  full  temperature  of  the 
last  retorts.  Sometimes  waste  heat  steam  boilers  are  located  behind  the 
furnaces  in  order  to  recover  this  extra  heat  in  a  useful  form,  but  as  the 
amount  of  power  required  around  a  zinc  plant  is  comparatively  small,  only 


532 


SMELTING  ZINC  ORES 


a  portion  of  the  waste  is  so  recovered.  For  that  reason  and  because  of  the 
increasing  cost  of  coal,  regenerative  furnaces  are  coming  into  increasing 
use.  In  these  furnaces  the  waste  heat  of  the  outgoing  gases  is  employed  to 
heat  brick  checkers,  which  in  turn  give  up  this  stored  heat  to  the  incoming 
air  and  gas.  In  that  way  a  saving  of  fully  60  per  cent  of  the  coal  consumed 
is  made,  although  at  the  cost  of  increased  labor  to  some  extent. 


FIG.  290. — Smelting  Furnace  in  Operation. 

One  of  these  regenerative  or  Seimens'  type  furnaces  is  shown  in  cA)ss- 
section  hi  Fig.  291,  and  Fig.  292  also  gives  in  outline  the  general  method 
of  admission  of  gas  and  air  through  the  regenerative  chambers  into  the 
laboratory  of  the  furnace,  and  their  movement  through  the  other  chambers 
and  flues  to  the  stack.  Periodical  reversals  of  the  gas  and  air  take  place 
at  intervals  of  half  to  one  hour  by  means  of  the  valves  shown,  whereby  the 
checkerwork-filled  chambers  are  alternately  heated,  and  again  give  up  their 
heat  to  the  incoming  gases. 


ZINC-SMELTING  PLANT 


533 


There  are  various  designs  of  these  furnaces,  but  nearly  all  work  on  the 
reversing  principle  as  above  outlined. 

Fig.  293  shows  in  detail  a  zinc-retort   in  place  in  the  furnace.     It  is 


iii'in    in'ii 


Air  Admission 
Valve 
Regenerative     A 
Checker-work_fef_ 
Floor 


SECTION  A-A 

FKJ.  1?91  — Cross-section  of  Zinc-smelting  Plant. 


FIG.  292. — Plan  of  Zinc-smelting  Plant. 

made  of  fireclay,  4  ft.  long  by  8.5  in.  diameter  and  with  walls  1.25  in. 'thick. 
In  the  figure  a  is  the  retort,  which  rests  on  a  ledge  on  the  rear  wall  of  the 
furnace  and  extends  just  through  the  thin  (4j-in.)  front  wall.  The  wall 
is  held  by  buck-staves  c  which  carry  the  tiles  upon  which  the  retorts  rest. 


534 


SMELTING  ZINC  ORES 


FIG.  293. — Zinc-smelting 
Retorts. 


The  whole  is  firmly  bound  together  with  tie-rods.     When  the  retort  has 
been  charged,  the  clay  condenser  b  is  set  in  place,  in  which  the  zinc  vapor 

issuing  from  the  retort  is  to  condense.  As 
seen  in  the  front  view,  the  space  between  two 
buck-staves  is  divided  by  shelves  which  form 
"  pigeon-holes,"  each  of  which  contains  two 
retorts.  The  retorts  having  been  set  in  place, 
the  opening  around  them  is  bricked  up  with 
pieces  of  brick  and  with  clay.  When  a  retort 
becomes  cracked  or  otherwise  useless,  it  can  be 
readily  removed  by  breaking  away  the  tem- 
porary wall,  and  another  retort  can  be  set  in  the  place  without  disturb- 
ing the  adjacent  retorts. 

OPERATING  THE  FURNACE 

The  roasted  blende,  or  oxidized  ore,  or  a  mixture  of  the  two,  is  thor- 
oughly mixed  with  fine  coal,  and  moistened  with  water  so  that  when 
thrown  into  the  retorts  it  will  pack  closely.  The  coal  used  for  reduction 
is  generally  a  low  volatile,  low  sulphur  coal,  preferably  anthracite.  In 
the  western  field  various  coals  are  used  for  this  purpose.  Sometimes 
anthracite  is  used  alone,  sometimes  what  is  known  as  "  dead  coal,"  a  non- 
coking  weathered  coal  found  near  the  surface  in  Kansas;  and  sometimes 


FIG.  294. — Charge-scoop. 

crushed  coke  or  coke  braize  is  employed  with  either,  or  even  a  mixture  of 
all  three.  The  amount  employed  is  generally  between  40  and  60  per  cent, 
dependent  on  the  character  and  grade  of  the  ore.  This  fuel  is  always  used 
crushed  not  coarser  than  1  in.,  but  generally  as  fine  as  \  in.  at  least. 

The  amount  of  ore  charged  per  retort  is  usually  about  60  Ib.  with  a 
proper  quantity  of  fuel.  The  amount  of  charge  for  a  furnace  is  placed  in 
one  or  more  cars  on  the  tracks  shown  in  front  of  the  furnace,  Fig.  288  and 
Fig.  289,  and  is  shoveled  directly  by  means  of  the  scoop,  Fig.  294,  into  the 
retort.  After  filling  the  retort  an  iron  rod  J-in.  thick  is  run  along  the  top 
of  the 'charge  next  the  retort  to  provide  for  a  vent  for  the  moisture  and 
gases  of  the  charge. 

As  the  retorts  are  filled  the  condensers  are  set  in  place  resting  on  sup- 
ports on  the  plates  in  front  of  the  furnace.  The  condenser  is  then  luted 


OPERATING  THE  ZINC  FURNACE  535 

or  loamed  around  the  joint  with  the  retort  by  a  finely  ground  and  damp- 
ened mixture  of  coal  and  field  loam;  hence  the  term  "  learning."  The 
open  end  is  loosely  filled  with  a  handful  of  a  mixture  of  coal  and  charge  or 
waste  material,  in  such  a  way  as  to  prevent  the  flowing  out  of  the  zinc,  yet 
permit  the  escape  of  the  reduction  gases. 

After  the  ore  is  charged  the  heat  of  the  furnace  is  gradually  raised  so  as 
to  drive  off,  first  the  water  added  to  moisten  the  charge,  then  the  volatile 
matter  of  the  coal,  and  the  carbonic  acid,  if  any  in  the  ore.  Finally  after 
about  two  or  three  hours  it  is  raised  to  the  heat  of  reduction  of  the  zinc. 

The  penetration  of  the  heat  from  the  outside  of  the  retort,  to  the 
inside  of  the  charge  being  progressive,  of  course  these  periods  of  the  process 
necessarily  overlap  each  other,  so  that  before  the  center  of  the  charge  is 
fully  dried  out  the  gases  are  coming  off  that  part  of  the  charge  in  contact 
with  the  walls;  and  before  these  latter  are  completely  expelled  metallic 
zinc  is  being  given  off.  This  causes  a  dilution  of  the  zinc  vapor  in  the 
earlier  stages  so  that  then  the  condensation  of  the  metal  is  incomplete,  and 
much  blue  powder  is  formed  as  well  as  zinc  vapor  lost.  From  this  period 
the  temperature  of  the  furnaces  outside  the  retorts  is  maintained  at  not  less 
than  1200°  C.,  increasing  to  1400°  C.  towards  the  finish  of  the  operation. 

After  the  charge  has  been  in  the  furnace  for  ten  or  twelve  hours  there  is 
generally  sufficient  metal  in  the  furnace  for  first  metal  draining.  The 
metal  drawer  brings  underneath  the  outlet  of  the  condenser  a  cast-iron 
ladle  swinging  on  a  crane,  and  by  means  of  a  special  tool  breaks  out  the 
stuffing  of  material  in  the  mouth  of  the  condenser.  Part  of  the  metal  runs 
out,  and  the  balance,  with  more  or  less  oxide,  blue  powder  and  portions  of 
the  charge  which  are  carried  out,  is  scraped  into  the  ladle  by  the  same  tool. 
This  tool  is  a  cast-iron  button  about  2  in.  diameter  riveted  on  the  end 
of  o-in.  iron  handle.  The  metal  is  poured  from  the  ladle  into  cast-iron 
molds.  The  resulting  slabs  are  about  12  by  18  by  1J  in.  and  weigh  about 
60  Ib.  each.  After  the  metal  is  drawn  the  condensers  are  closely  stuffed 
as  before,  and  the  operation  continues  without  interruption.  At  the  end 
of  about  eight  hours  more  the  furnace  is  drawn,  and  again  at  the  end  of 
about  four  hours,  when  the  operation  is  finished. 

The  condensers  are  then  chiseled  loose  from  the  retorts  and  moved  to 
one  side.  The  loose  unworked  charge  and  oxide  around  the  mouth  of 
the  retorts  are  scraped  out,  and  the  furnace  plates  and  floor  are  thoroughly 
cleaned  up.  The  cleanings,  together  with  the  oxides,  skimmings,  etc., 
made  during  the  process  are  put  aside  to  be  charged  again,  and  the  furnace 
is  ready  for  cleaning  out.  ' 

The  covers  of  the  openings  in  front  of  the  furnace  are  now  removed, 
and  the  operatives  scrape  out  the  residues.  These  residues  flow  through 
the  openings  into  the  cellar  below,  where  generally  cars  are  ready  to  receive 
them.  Sometimes  other  methods  are  used  to  remove  the  bulk  of  the  res- 


536  SMELTING  ZINC  ORES 

idues  from  the  retorts,  but  scrapers,  or  bumper^,  so  called,  have  to  be 
employed  for  a  final  cleaning.  The  castings  and  floor  in  front  of  the  fur- 
nace are  now  swept  clean  of  the  spent  residues,  the  covers  of  the  cellar 
holes  are  replaced  and  the  furnace  is  ready  to  recharge.  The  operation 
from  this  point  is  as  described  above. 

After  the  furnace  is  cleaned  out  any  broken  or  corroded  retorts  are 
readily  discovered  and  are  removed.  This  is  done,  and  they  are  replaced 
by  new  retorts  which  have  been  brought  up  to  red  heat  in  a  kiln  for  that 
purpose,  and  without  being  cooled  down  are  pushed  into  place  in  the  hot 
furnace.  When  the  retorts  are  in  place  they  are  closed  into  the  furnace 
by  a  fireclay  partition  which  fits  closely  around  the  mouth  and  closes  the 
opening  to  the  furnace  so  as  to  retain  the  fire. 

The  heat  of  the  furnace  is  allowed  to  fall  slightly  during  the  time  of 
charging  and  changing  broken  retorts,  but  at  no  time  during  the  cam- 
paign, which  may  last  five  to  seven  years,  is  the  furnace  allowed  to  cool 
off. 

MANUFACTURE   OF  RETORTS  AND  CONDENSERS 

Retorts  to  withstand  the  high  temperature  and  corrosive  action  of 
the  charge  are  made  of  the  most  compact  and  durable  material.  The 
material  consists  of  a  mixture  of  "  chamotte,"  "  grog,"  or  "  cement,"  of 
burned  fireclay,  firebrick,  or  tile  free  from  slag.  It  is  ground  to  about 
6-mesh  size  and  mixed  with  an  approximately  equal  amount  of  raw  fireclay. 
The  mixing  is  done  in  a  pug-mill,  using  water  to  form  a  stiff  mud,  which  is 
allowed  to  stand  some  tune  covered  with  wet  sacking  to  season  and  to 
develop  the  plasticity.  It  is  again  put  through  the  pug-mill,  and  finally 
made  into  retorts  in  a  hydraulic  retort-making  machine  under  the  pressure 
of  about  3000  Ib.  per  square  inch.  A  machine  of  this  kind  makes  twenty 
retorts  or  more  per  hour. 

The  success  of  the  retorting  operation  depends  upon  the  durability  of 
the  retorts,  and  for  this  reason  a  careful  selection  of  the  clay  is  the  first 
necessity.  With  one  or  two  exceptions,  all  the  zinc  smelters  in  the  United 
States  use  a  fireclay  found  in  large  quantities  in  the  Mississippi  Valley, 
chiefly  at  or  near  St.  Louis.  It  is  probable  that  the  good  results  obtained 
with  retorts  made  from  this  clay  are  partly  the  result  of  a  familiarity  and 
knowledge  of  its  qualities  for  the  purpose,  as  the  attempts  mad^  to  use 
other  clays  of  apparently  superior  physical  and  chemical  properties,  have 
for  the  most  part  resulted  in  failure.  The  analysis  of  this  clay,  generally 
known  as  "  Cheltenham  "  clay,  is  about  as  follows,  on  an  air-dried  sample: 

Per  Cent. 
A12O3 30-33 

SiO2 50-45 

Bases 4-5 

and  loss  on  ignition,  including  water  and  organic  matter  about  15  per  cent. 


RETORT  MAKING  537 

Condensers. — Condensers  are  made  of  less  refractory  clay  than  retorts. 
They  are  not  subjected  to  high  temperature,  but  must  withstand  much 
handling  and  severe  treatment.  They  last  eight  to  twelve  days,  and  cost 
3  to  4  cents  each. 

Drying  the  Retorts. — The  finished  retorts  as  they  are  removed  from  the 
machine  are  placed  in  vaults  or  compartments  holding  500-1000  retorts, 
according  to  size  of  works.  These  vaults  are  provided  with  steam  coils 
beneath  the  floor,  for  heating  and  drying.  In  general  the  temperature  of 
the  vault  is  kept  low  during  the  first  week  or  more,  to  allow  slow  evapora- 
tion of  the  added  water,  then  is  gradually  raised  during  successive  periods 
until  the  retort  is  thoroughly  freed  from  moisture.  The  period  of  season- 
ing may  hardly  be  less  than  four  weeks,  and  may  better  be  prolonged  for 
at  least  twelve  weeks.  It  is  considered  better  that  it  should  be  a  slow  opera- 
tion, but  requirements  in  this  respect  differ  with  different  clays. 

Annealing. — After  the  retorts  are  thoroughly  seasoned  and  dried  they 
are  available  for  use  in  the  furnaces  as  required.  Each  day  the  number  of 
retorts  experience  has  shown  to  be  required  are  placed  in  a  small  furnace 
or  "  temper-kiln  "  cold,  and  the  heat  gradually  raised  so  that  first  the  com- 
bined water  is  removed  and  later  the  heat  of  the  kiln  is  brought  up  to  full 
redness.  About  twenty-four  hours  is  usually  required  for  this  purpose, 
and  at  the  proper  time,  as  required  in  the  operation  of  the  furnace,  these 
retorts  are  taken  while  still  red  hot  and  placed  in  position  in  the  smelting 
furnace. 

LOSS  IN  THE  PROCESS 

The  losses  in  smelting  of  zinc  occur,  as  may  be  expected,  in  every  process 
to  which  the  ore  is  subjected. 

In  blende-roasting  there  is  a  mechanical  loss  from  spilling  of  ore,  and 
from  flue  dust,  and  a  metallurgical  loss  of  zinc  in  the  form  of  fume  due  to 
the  volatilization  of  zinc. 

In  smelting  the  roasted  blende  or  carbonate  silicate  or  oxide,  there  are 
again  mechanical  losses  in  spilling  and  dust,  but  the  more  serious  losses 
are  those  of  the  metallurgical  process  itself.  These  may  be  summed  up  as : 
(1)  Infiltration  of  zinc  through  the  retort  and  absorption  of  zinc  by  the 
retort  itself;  (2)  loss  in  fume  from  the  condenser  due  to  uncondensed 
zinc;  (3)  fume  or  zinc  vapor  remaining  in  the  retort  at  the  conclusion  of 
the  process;  (4)  zinc  remaining  in  the  residuum  after  removal  from  the 
retort. 

In  the  roasting  of  blende  the  losses  will  amount  to  at  least  1  per  cent 
and  will  average  2.  In  the  case  of  fine  ore,  dust  losses  may  increase  this 
proportionally. 

In  retorting  roasted  blende  which  averages  40  per  cent  zinc  before 
roasting,  the  losses  will  average  16  per  cent;  where  the  zinc  tenor  is  as 


538  SMELTING  ZINC  ORES 

high  as  50  per  cent  the  loss  will  approximate  13  per  cent  and  when  60 
per  cent  material  is  handled  the  average  loss  will  approximate  10  to  11 
per  cent.  Based  on  the  original  unroasted  material,  the  total  of  all  losses 
will  be  approximately  for  40  per  cent  material  18  per  cent,  50  per  cent  will 
be  15  per  cent,  and  60  per  cent  about  12  to  13  per  cent.  Special  conditions 
and  special  characteristics  of  ores  will  of  course  modify  these  figures. 

In  smelting  carbonate  and  silicate  ores,  which  average  lower  in  zinc 
content,  the  per  cent  of  loss  will  be  less.  These  ores  for  the  most  part  have 
zinc  contents  of  30  to  40  per  cent  and  the  zinc  loss  will  be  from  12  per  cent 
for  the  better  grades  to  17|  or  even  20  per  cent  for  the  poorer. 

COST  OF  SMELTING 

The  cost  of  the  smelting  of  blende  of  course  varies  much  with  different 
localities.  The  large  natural  gas  fields  found  in  the  Kansas-Oklahoma 
territory  enabled  cheap  operations  to  be  carried  on  at  points  so  situated 
with  respect  to  the  most  important  zinc  deposits  of  the  country  that  the 
freight  charges  on  the  ore  to  the  works  and  on  the  metal  to  the  point  of 
consumption,  were  reduced  to  a  minimum.  The  costs  in  that  field,  there- 
fore, have  been  and  are  to-day  exceedingly  favorable,  considering  the  char- 
acter of  the  operation.  Exclusive  of  gas  cost  these  charges  in  1912  are 
about  as  follows: 

Per  Ton  Blende. 

Unloading,  crushing,  drying,  and  sampling $0 . 35 

Roasting,  labor  and  repairs 0 . 50 

Smelting.  Per  Ton  Roasted. 

Labor,  direct $4 . 00 

Reduction  fuel 1 . 80 

Retorts,  condensers,  etc 0 . 60 

Repairs,  maintenance,  power,  etc 0 . 75 

Charging,  etc 0 . 35 

General  labor 0 . 20 

Superintedance,  office  expenses,  etc 0 . 75 


$8.45 
Roasting  loss  13  per  cent =     1 . 10  7.35 


Total,  exclusive  of  gas $8 .  S$) 

The  cost  of  gas  varies  with  the  conditions  surrounding  the  plant. 
In  the  early  days  of  any  gas  field  the  cost  is  apparently  very  low,  but 
increases  rapidly  with  the  exhaustion  of  neighboring  fields.  Probably 
a  fair  average  cost  of  gas  per  ton  of  blende  for  both  roasting  and  smelting 
would  be  $2  to  $2.25  per  ton.  This  added  to  the  other  costs  makes  the 
total  cost  average  from  $10.25  to  $10.50  per  ton  for  the  whole  operation. 

In  the  coal  fields  the  cost  of  fuel  is  higher,  but  owing  to  the  location 


PRICES  OF  ZINC  ORES  539 

generally  chosen  for  these  plants  using  coal  for  fuel,  there  is  a  good  market 
for  sulphuric  acid  at  such  points,  and  the  manufacture  of  acid  as  a  by- 
product reduces  this  cost  to  a  point  generally  much  below  the  cost  at 
natural  gas-using  plants. 

PRICE   OF  ZINC  ORES  AND  SMELTERS  IN  1919 

Mississippi  Valley  ores  are  bought  as  at  Joplin,  Mo.,  at  a  quoted 
price,  on  a  basis  of  60  per  cent  zinc  contents.  As  this  varies  up  or  down, 
$1  per  unit  is  added  or  subtracted  from  the  base  price.  Calomine  is 
bought  on  a  40  per  cent  basis.  We  may  quote  for  a  certain  blende  $47, 
and  for  calamine  $30  per  short  ton. 

Western  Zinc  Ores. — On  a  40  per  cent  basis  we  may  have,  for  example, 
$20  per  ton  paid  for  a  calamine  ore  or  $13.50  for  a  sulphide.  A  variation  of 
$1  per  ton  per  unit  is  made  up  or  down.  Besides  this  65  per  cent  of  the 
zinc  contents  is  added  or  subtracted  as  the  market  price  of  zinc  rises  or  falls. 

Iron  over  2  per  cent  is  penalized  at  $1.50  a  unit;  arsenic  or  fluorine 
are  not  permitted.  The  gold  and  silver  may  be  paid  for  at  65  per  cent  of 
their  market  value.  The  above  figures  are  based  on  zinc  at  8  cents  per 
pound. 

Zinc. — Quotations  in  the  United  States  in  1919  are  given  in  cents  per 
pound,  thus:  4.60  cents,  St.  Louis;  4.75  to  4.80  cents,  New  York.  The 
London  market  is  quoted  at  9.15  shillings  for  good  ordinaries  (ordinary 
brands)  and  20  shillings  for  specials  (the  purer  zinc).  St.  Louis  is  near 
the  zinc-producing  district  of  Kansas,  Missouri,  and  Illinois,  and  hence 
has  a  lower  price  for  spelter  than  New  York. 

The  European  Price  of  Zinc  Ores. — The  value  of  a  zinc  ore  depends 
upon  its  content  in  zinc  and  the  absence  of  objectionable  impurities, 
such  as  iron,  manganese  and  lime,  which  form  fusible  slags,  and  increase 
the  corrosion  of  the  retorts,  or  of  lead,  cadmium,  arsenic  and  antimony, 
which  contaminate  the  spelter  and  so  lower  its  market  value. 

For  many  years  large  quantities  of  zinc  ores  have  been  sent  to  Ant- 
werp, Belgium,  and  to  Swansea,  Wales,  from  Sardinia,  Algeria,  and 
Spain,  and  of  late  from  Colorado,  via  the  Gulf  ports  of  Galveston  and  New 
Orleans. 

Based  on  ores  of  46  per  cent  zinc  and  upward  the  price  of  the  ore,  with 
costs,  insurance  and  freight  paid,  was  the  London  price  less  8  units  and  95 
per  cent  of  the  assay  value,  and  less  a  returning  or  treatment-charge  of 
perhaps  £33  per  ton. 


CHAPTER  XLIII 


ZINC  REFINING 

Refining  is  for  the  purpose  of  converting  low-grade  spelter  into  grades 
suitable  for  high-grade  brass.  For  such  grades  from  5  to  15  cents  per 
pound  over  prime  Western  spelter  was  paid  during  the  Great  War. 

GRADES  OF  ZINC 

Electrolytic  spelter  of  99.9  to  99.95  per  cent  is  the  purest  made. 
Aside  from  this  the  great  bulk  of  spelter  made  by  distillation,  and  called 
virgin  spelter,  has  been  divided  into  four  grades,  as  follows: 


Grade. 

Pb, 
Per  Cent. 

Fe, 
Per  Cent. 

Cd, 
Per  Cent. 

A    High  ,<.. 

0.07 

0.03 

0.05 

B    Intermediate                                

0  20 

0   03 

0.50 

C    Brass  special 

0  75 

0  04 

0  75 

D    Prime  Western  

1.50 

0.08 

In  distillation  lead  is  volatilized  and  carried  over  into  the  spelter. 
For  the  grade  "  brass  special  "  the  lead  in  the  ore  should  be  below  1  per 
cent.  Iron  makes  less  trouble,  and  may  occur  in  ores  up  to  10  to  12  per 
cent.  Cadmium  is  even  more  easily  distilled  than  zinc  and  so  first-draw 
zinc  contains  the  most  of  it.  It  is  not  considered  to  be  detrimental  in 
small  proportions. 

Redistilling. — To  prepare  for  this  the  ordinary  ore-smelting  furnace 
has  its  lower  row  of  retorts  removed  and  the  butts  of  the  upper  rows  are 
placed  upon  the  shelves  of  the  next  lower  row.  This  gives  the  retort  an 
inclination  of  8  to  10  in.  in  its  length  and  permits  it  to  hold  a  bath  01  molten 
spelter.  Since  in  redistilling  a  lower  temperature  (950°  C.)  can  be  carried 
than  in  smelting,  the  flue  checkers  are  opened  and  the  combustion  gas  is 
burned  under  natural  draft.  There  results  a  thin  flame,  a  more  uniform 
heat  and  better  operating  conditions.  The  condenser  used  may  be  an 
ordinary  one,  having  a  dam  block  in  half  the  larger  end ;  or  again  a  hand- 
made cone  condenser  may  be  preferred.  This  smaller  end,  7  in.  diameter 
with  a  tile  dam,  is  luted  into  the  retort,  while  the  larger  end,  10  in.  diameter 

540 


ZINC  REFINING  541 

is  closed  with  a  fireclay  plate  with  tap-hole  openings,  top  and  bottom. 
These  openings  are  stuffed  up  during  distillation.  The  low-grade  spelter 
which  is  to  be  refined  is  cast  into  sticks  20  in.  long  by  1^  in.  square.  Four 
or  five  of  these  make  a  charge  for  one  retort,  these  being  charged  imme- 
diately after  each  drawing.  The  sticks  are  inserted  by  the  top  opening. 
This  is  stuffed,  and  in  five  or  ten  minutes  the  sticks  are  melted  and  in 
process  of  distillation.  Drawing  is  done  every  six  hours.  To  do  this  the 
top  opening  is  spiessed  or  pricked  open,  in  order  to  relieve  the  internal  gas 
pressure,  the  scratcher  is  inserted  in  the  bottom  hole  and  the  contained 
metal  and  blue  powder  drawn  into  the  ladle.  The  redistilled  spelter  is 
cast  into  plates  and  taken  to  a  reverberatory  equalizing  furnace  for  recast- 
ing into  plates  of  uniform  high-grade  spelter. 

As  the  bath  of  metal  in  the  retort  becomes  enriched  in  lead  and  iron  it 
must  be  removed.  This  is  done  by  omitting  the  charge  for  say,  twenty- 
four  hours  on  one  section  of  nine  retorts,  and  the  next  day  taking  down  the 
condenser  of  this  section  and  scraping  out  the  metal  called  "  bottoms," 
using  a  large  scraper.  In  case  of  a  leaky  retort  all  is  removed  and  the 
retort  at  once  replaced.  The  leady  bottoms  drawn  into  a  ladle  are  cast 
into  plates  and  taken  to  a  remelting  furnace  where  any  excess  zinc  is  sep- 
arated. The  lead,  tapped  from  this  furnace  and  carrying  1  to  2  per  cent 
zinc,  is  sold  to  lead  refineries.  The  blue  powder,  skimmings,  etc.,  from  the 
redistilling,  remelting  and  equalizing  furnaces  are  returned  to  the  ore 
furnaces  for  resmelting.  All  scrap  metal  is  sent  to  the  remelting  furnace  to 
be  recast  into  sticks  of  recharging.  Five  men  operate  the  two  furnaces  of 
the  block. 

To  refine  spelter,  a  furnace  resembling  that  shown  in  Fig.  241  is  used, 
but  the  reverberatory  firebox  is  in  two  parts,  and  provision  is  made  to 
charge  the  spelter  close  to  the  bridge.  It  holds  30  tons  of  spelter  when 
full,  and  in  it  10  tons  can  be  refined  in  twenty-four  hours.  The  metals 
separate  into  layers.  At  the  bottom  is  the  lead;  the  iron  forms  with  the 
zinc  and  part  of  the  lead,  a  difficultly  fusible  alloy  that  floats  on  the  lead, 
and  uppermost  is  the  stratum  of  pure  zinc.  By  means  of  an  iron  rod  in- 
serted into  the  bath,  layers  are  distinguished,  the  zinc  being  soft,  the  iron- 
lead-zinc  alloy  (called  "  hard  zinc  ")  being  mushy,  and  the  molten  lead  at 
the  bottom  soft.  The  underlying  lead  is  removed  weekly.  A  cylinder  or 
pipe  closed  at  the  lower  end  is  sunk  below  the  lead  layer.  The  plug  is 
then  knocked  out  and  the  lead,  rising  in  the  cylinder,  is  ladled  into  molds. 
The  zinc  of  the  top  layer  is  ladled  out  daily  into  molds,  and  it  retains  1  to 
1.25  per  cent  lead.  The  hard  zinc  layer  is  removed  when  opportunity 
offers.  To  do  this  the  zinc  is  ladled  out  first,  the  lead  is  next  removed, 
and  finally  the  mushy  mass  of  ferruginous  metal  is  removed  with  ladles 
perforated  so  that  the  lead  drains  off.  This  hard  zinc  is  sold  for  the  manu- 
facture of  Delta  or  Sterro-metal. 


542  ZINC  REFINING 


THE  DE  SAULLES  REDISTILLATION  METHOD 

A  furnace  is  used  like  one  side  of  an  ordinary  furnace  block.  The  retorts 
are  inclined  7  in.  in  their  length  and  extend  4  to  5  in.  through  the  back  wall. 
At  the  top  of  the  protruding  back  is  an  opening  for  charging  the  retort  with 
molten  spelter.  At  the  bottom  is  a  small  tap-hole  for  removal  of  the  leady 
bottoms.  Both  openings  are  tightly  closed  with  clay  except  when  charging 
or  tapping.  There  is  an  ordinary  condenser  properly  clayed  to  the  retort. 
Each  retort-distilling  furnace  of,  say,  200  retorts,  is  served  by  a  25-ton 
remelting  furnace  and  a  25-ton  equalizing  furnace. 

Low-grade  spelter  for  redistillation  consists  of  second  or  third-draw 
metal,  and  contains  1.5  to  3.0  per  cent  lead,  0.03  to  1.0  per  cent  iron  and 
0.03  to  0.07  cadmium.  The  redistilled  spelter  will  average  0.10  per  cent 
lead,  0.01  per  cent  iron  and  0.04  per  cent  cadmium,  a  better  grade  than 
"  intermediate,"  as  given  in  the  table. 

REFINING  SPELTER  WITHOUT  REDISTILLATION 

The  principle  of  this  method  consists  in  remelting  low-grade  spelter  in 
a  reverberatory  furnace  with  a  reducing  flame,  and  letting  the  molten 
bath  stand  until  the  metal  separates  into  layers  according  to  the  specific 
gravity  of  the  different  metals,  the  lower  part  of  the  bath  consisting  of  a 
leady  zinc  and  the  upper  part  of  spelter  nearly  free  from  lead.  The  lower 
layer  is  then  tapped,  or  removed  otherwise.  The  separation  or  refining 
must  be  done  at  a  temperature  near  the  melting-point  of  zinc,  since  the 
higher  the  temperature  the  more  persistently  does  the  zinc  retain  lead. 
Under  the  most  favorable  circumstances  the  lead  content  of  the  spelter 
is  reduced  to  1  to  1.25  per  cent. 

ELECTROLYTIC  ZINC 

Due  to  the  demand  for  a  pure  metal  for  the  making  of  cartridge  brass 
during  the  great  war,  electrolytic  zinc  was  held  at  a  premium  of  2  or  3 
cents  per  pound  over  the  then  high  price  of  other  grades.  This  gave  a 
great  impetus  to  developing  a  successful  electrolytic  method  for  its  manu- 
facture. 

The  process  consists  briefly  in  roasting  the  ore,  dissolving  in  dilute 
sulphuric  acid,  filtering  the  solution,  precipitating  the  other  bases  by  means 
of  zinc-dust,  filtering  to  obtain  a  solution  containing  zinc  only,  and  pre- 
cipitating this  in  metallic  form  by  electrolysis. 

Roasting  of  Zinc  Sulphide  to  Oxide  and  Sulphate. — Zinc  sulphate  can 
be  formed  through  any  one  of  the  following  reactions : 


ELECTROLYTIC  ZINC  543 

(9)  ZnS+4O  =  ZnSO4. 

(10)  ZnS+3O  =  ZnO+SO2. 

(11)  ZnO+SO2+O  =  ZnSO4  (a). 

(12)  ZnO+SO2+Fe2O3  =  ZnSO4+2FeO  (6). 

(13)  2FeO+O  =  Fe203. 

(14)  ZnO+SO3  =  ZnSO4  (c) . 

There  is  considerable  evidence  that  the  first  reaction  is  responsible  for 
most  of  the  sulphate  formed.  The  only  gaseous  reagent  is  oxygen  and 
there  are  no  gaseous  reaction  products,  therefore,  the  oxygen  concentration 
alone  should  mainly  determine  the  amount  of  sulphate  formed.  Reac- 
tions (11)  and  (12)  involve  two  gaseous  reagents,  so  that  the  amount  of 
sulphate  formed  will  be  determined  mainly  by  the  product  of  the  concen- 
trations of  oxygen  and  sulphur  dioxide. 

The  ore,  of  30-mesh  size,  is  a  blende  concentrate  of  25  per  cent  sulphur. 
It  is  roasted  in  a  Wedge  roaster,  being  given  a  close  oxidizing  roast  that 
brings  it  down  to  2  per  cent  sulphur,  yielding  a  product  of  zinc  oxide  and 
sulphate.  This,  after  cooling,  is  stored  in  feed-bins,  whence  it  is  drawn 
off  into  agitating  vats.  Here  it  is  treated  for  twenty-four  hours  with  dilute 
sulphuric  acid.  The  acid  dissolves  the  zinc  oxide  and  small  amounts  of 
copper  and  cadmium  also  present.  The  pulp  is  now  passed  over  a 
classifier  which  removes  the  sand,  the  slime  then  going  to  an  Oliver  filter. 
Both  the  sand  and  the  slime  are  stored  and  are  sold  at  a  profit  to  the 
smelter  as  containing  silver  and  gold  and  a  little  lead. 

There  remains  the  clear  solution,  which  is  passed  to  another  agitator. 
Here  it  is  treated  to  a  small  addition  of  granulated  zinc  made  in  the  melting 
house  from  zinc  produced  by  electrolysis.  After  a  prolonged  treatment,  in 
which  the  zinc  dust  precipitates  all  the  other  bases,  the  product  is  pumped 
through  a  closed  Sweetland  filter,  where  the  zinc-dust  and  the  bases  are 
removed,  leaving  a  clear  solution  containing  zinc  only.  This  solution  is 
stored  hi  vats  and  is  drawn  off  as  needed  to  the  electrolytic  tanks  of  the 
tank-house,  arranged  and  operated  precisely  like  the  tank-house  of  copper 
refining.  Each  tank  contains  eighteen  anodes  and  nineteen  cathodes,  the 
current  being  hi  parallel  with  a  tank-resistance  of  0.4  volt.  The  anodes 
are  of  lead,  21  by  36  in.,  while  the  cathodes  are  of  aluminum  24  by  36  in. 

When  the  zinc  has  been  deposited  to  the  depth  of  J  in.  on  the  cathodes 
these  are  removed  from  the  tank,  washed,  and  the  zinc  is  stripped.  This 
is  melted  in  a  reverberatory  furnace,  and  melted  into  commercial  plates, 
10  X 16X2  in.  A  small  portion  of  the  molten  metal  is  granulated,  it  being 
placed  in  a  small  reservoir  whence  it  runs  in  a  stream  the  size  of  a  knitting 
needle  so  as  to  be  caught  by  a  horizontal  air  jet.  This  blows  it  to  powder 
and  it  is  caught  within  a  sheet  steel  bin  or  chamber. 


PART  VIII 
PLANT,   EQUIPMENT  AND  THEIR  COSTS 


CHAPTER  XLIV 
LOCATION,  EQUIPMENT  AND  ERECTION 

LOCATION  OF  WORKS 

A  mine  has  practical  value  only  when  permanence  is  reasonably 
assured.  Then,  a  single  unit  of  a  mill  may  be  erected  where  the  process 
that  has  been  chosen  may  be  perfected ;  after  that  other  units  can  be  added. 

Mines  Works. — If  a  mining  company  builds  a  mill  for  the  treatment 
of  its  own  ore,  it  is  usual,  to  save  freight  or  hauling  expense,  to  place  the 
plant  as  close  to  the  mine  as  the  securing  of  a  suitable  site  and  water-supply 
permits.  In  case  of  a  smelting  works  where  flux  and  fuel  is  to  be  brought 
in  and  a  heavy  product  shipped,  then,  hi  addition  to  a  good  site  and 
water  supply,  nearness  to  a  railroad  is  to  be  considered. 

Custom  Works. — This  can  judiciously  buy  ores  from  neighboring  mines, 
but  needs  a  site  convenient  to  the  chief  source  of  supply  and  to  the  coke 
and  fuel  that  it  must  use.  For  such  a  plant  a  point  should  be  chosen 
where  several  railroads  give  rise  to  competition  in  freight  rates  and  where 
labor  is  abundant.  Low  freight  rates,  abundant  labor,  and  low  money 
rates  combine  in  locating  a  custom  works. 

Iron  and  Steel  Plants. — Thus  iron  and  steel  manufacture  has  centered 
about  Pittsburg,  Pa.,  because  coke,  coal,  and  natural  gas  are  abundant, 
and  because  a  good  market  is  found  there  for  the  products.  On  the  other 
hand,  the  iron  smelter  at  Pittsburg  must  pay  for  freight  from  mine  to 
furnace,  $2.25  per  ton,  and  must  carry  a  large  supply  of  ore  to  last  through 
the  winter  months  when  navigation  is  closed.  The  United  States  Steel 
Corporation,  the  largest  manufacturer  of  iron  and  steel  in  the  world,  has 
erected  a  plant  near  the  iron  ranges  at  Duluth,  Minn.,  for  reasons  shown 
below. 

Effect  of  Water  Carriage. — Vessels  carrying  iron  ore  to  Lake  Erie 
ports  can  return  with  cargoes  of  coke  or  coal  to  supply  the  Duluth  furnaces, 
which  then  have  a  local  market  for  their  pig-iron,  and  do  not  need  to 
"  stock  up  "  with  a  winter's  supply  of  fuel.  Figuring  roughly  that  2f  tons 
coal,  made  into  coke  and  into  producer-gas,  is  required  to  make  a  ton  of 
steel,  there  is  a  slight  advantage,  as  to  fuel,  in  making  coke  in  by-product 
ovens  at  Duluth,  Minn.,  and  using  gas-engines  which  utilize  the  blast- 
furnace gases  to  the  best  advantage.  Nearly  two  tons  of  iron  ore  must  be 
sent  to  Eastern  furnaces  to  produce  this  one  ton  of  steel. 

547 


548  LOCATION,  EQUIPMENT  AND  ERECTION 

Zinc  Works. — It  is  seen  from  the  cost  of  producing  zinc,  that  3.5  tons  of 
coal  are  needed  per  ton  of  ore.  Thus  it  is  cheaper  to  convey  ore  to  fuel, 
than  coal  to  the  mine  where  ore  is  produced.  Near  Joplin,  Mo.,  there  is 
ore  and  also  fuel;  we  expect,  therefore,  to  find  the  zinc-smelting  works 
working  there  to  the  best  advantage.  The  region  is  made  more  favorable 
by  the  fact  that  natural  gas  is  to  be  had  there. 

Silver-lead  Works. — With  respect  to  silver-lead  works  using  lead 
as  a  collector  of  other  metals,  the  favored  places  have  been  found  to  be 
railroad  centers,  such  as  Denver,  Pueblo,  and  Salt  Lake.  From  12  to  15 
per  cent  coke  is  used  in  the  charge  in  smelting,  so  that  nearness  to  coal- 
fields is  not  the  all-important  condition.  On  the  other  hand,  ores  are 
available  there  in  proportion  favorable  to  combining  profitably  with  one 
another.  The  lead  of  one  ore  and  the  iron  of  another,  being  combined, 
serve  the  requirements  of  smelting. 

The  silver-lead  and  copper  custom  smelters  carry  a  supply  to  last  from 
two  to  four  weeks,  but  at  a  mines  works  provision  is  needed  but  for  one  to 
two  days'  running. 

Mills. — In  treating  ore  by  milling  and  cyaniding,  the  amount  of  fuel 
and  other  supplies  required  is  small,  and  hence  the  natural  place  for  the 
work  is  near  the  mine  that  produces  the  ore,  provided  the  extraction,  or 
recovery  of  the  precious  metals,  is  high.  When,  however,  the  ore  is  refrac- 
tory and  the  recovery  is  low,  it  pays  to  ship  the  ore  to  smelting  works  that 
guarantee  a  high  extraction. 

NATURE  OF  THE  SITE  TO  BE  CHOSEN 

Both  side-hill  and  flat  sites  are  chosen.  For  the  side-hill  or  terraced 
site,  much  used  for  mills,  the  ore  is  arranged  to  advance  or  flow  by  gravity 
from  one  operation  to  the  next,  using  elevators  less  than  on  a  level  site. 
Since  ore  and  mill  or  smelter  products  are  so  readily  moved  by  cars  and 
industrial  locomotives  which  easily  reach  the  higher  levels  of  the  works, 
certain  objections,  when  material  was  man-handled,  can  be  said  to  fall 
away.  On  the  other  hand  the  flat  site  has  these  advantages:  (1)  The 
first  cost  of  the  works  is  less,  since  heavy  grading  and  retaining  walls  are 
not  needed.  (2)  In  the  side-hill  works  the  different  parts  of  the  plant 
must  be  placed  in  a  definite  and  constrained  order  to  obtain  the  needed 
fall,  whereas  on  the  flat  site  one  can  expand  in  any  direction,  that  is,  the 
parts  that  have  to  be  far  apart  on  the  inclined  site,  can  be  placed  close 
together  on  a  flat  one.  A  building  on  a  flat  site  can  be  better  ventilated 
than  one  crowded  against  the  slope.  On  a  flat  site  elevators  are  more 
numerous,  but  they  are  also  convenient  for  delivering  just  where  you 
want  and  at  a  low  cost  per  ton. 

Iron  and  steel  plants,  the  largest  in  the  world,  are  constructed  on  level 
ground.  The  ore  is  unloaded  direct  from  vessels  to  the  stock  pile,  using 


PLANT  SITES  549 

grab-buckets  holding  5  to  10  tons  each.  If  transported  in  cars,  the  cars 
are  loaded  in  a  similar  manner.  The  custom  is  to  use  hopper-bottom  cars, 
from  which  the  ore  drops  into  the  charging  bins,  and  thence  by  charge-cars 
is  conveyed  to  the  furnace-skip.  By  the  skip  it  is  hoisted  100  ft.  to  the 
furnace-top.  Many  recent  silver-lead  smelting  plants  occupy  level  sites, 
but  the  dumping  ground  for  slag  is  at  a  lower  level. 

For  iron  works  little  attention  is  paid  to  the  location  of  the  slag-dump. 
There  is  no  hesitation  in  sending  the  slag,  if  necessary,  a  mile  away  by 
locomotive  to  be  dumped. 

On  the  other  hand  at  copper  and  lead  smelting  works  the  designer  likes 
to  have  at  least  two  levels.  At  the  largest  copper  reduction  plant  in  the 
world,  at  Anaconda,  the  side-hill  site  has  been  chosen.  Metallurgical 
mills  are  very  commonly  on  steep  side-hills. 

Mill-sites. — On  the  unclaimed  mineral  lands  of  the  Western  United 
States,  title  is  secured  from  the  general  Government  for  a  mill-site  for 
reduction  works,  five  acres  in  extent,  either  in  connection  with  a  mining 
claim  (on  a  theory  that  each  lode  claim  is  entitled  to  a  mill-site)  or  as  a  site 
for.  an  independent  or  custom  reduction  plant.  A  reduction  company, 
operating  a  mill,  must  dispose  of  the  tailing  it  produces,  and  of  the  water 
discharged,  not  encroaching  upon  the  property  of  other  people,  and  it  is 
responsible  for  all  damages.  A  company  must  not  let  tailing  that,  at  a 
reasonable  cost,  can  be  impounded,  flow  into  a  stream,  nor  run  into  waters 
where  liable  to  interfere  with  navigation.  The  right  or  custom  of  dumping 
on  the  valueless  land  of  lower  mining  claims  is  general, .  except  that  the 
practice  must  do  no  damage  to  the  property  of  owners  below. 

A  reduction  company  can  take  up  lands  for  a  ditch  or  flume  from  unap- 
propriated public  land,  and  the  claim  cannot  be  interfered  with  by  later 
locators;  but  the  owner  of  such  a  ditch  or  flume  is  responsible  for  damage 
arising  from  breaks  or  overflows. 

The  same  rule  holds  with  respect  to  roads  and  trails.  In  Colorado, 
mining  claims  are  subject  to  the  right-of-way  of  parties  hauling  ore  over 
them,  but  in  other  States  the  location  gives  exclusive  control,  except  that  a 
water,  electric,  or  railroad  company  can  take  it  under  the  law  of  eminent 
domain  by  giving  a  fair  compensation  for  it. 

Damage  from  Smoke. — The  smoke  from  the  smelting  works,  especially 
those  treating  sulphide  ore  in  quantity,  delivers  into  the  atmosphere  many 
tons  of  sulphur-fume  daily,  as  well  as  fine  flue-dust  carried  out  of  the  stack 
by  the  draft.  This  diffuses  through  the  atmosphere  and  is  carried  by  the 
wind  to  trees  and  the  crops  of  the  land.  If  not  diluted,  it  blights  vegeta- 
tion, and  naturally  the  farmers  organize  to  secure  damages,  or  to  close  the 
works.  The  question  of  what  to  do  to  avoid  the  difficulties  is  a  serious 
one,  and  to-day  when  pyrite  smelting  and  extensive  roasting  of  sulphide 
ores  is  carried  on,  the  trouble  can  not  be  altogether  overcome.  Thus  far 


550  LOCATION,  EQUIPMENT  AND  ERECTION 

the  solution  has  consisted  in  locating  the  works  in  places  where  there  is 
little  vegetation  to  be  damaged,  or  in  discharging  the  fume  into  the  atmos- ' 
phere  from  high  stacks.  It  may  be  said  that  the  latter  expedient  lessens 
but  does  not  altogether  obviate  the  difficulties.  The  metallurgist  must 
therefore  give  serious  consideration  to  the  matter,  otherwise,  after  erecting 
and  starting  the  operation  of  a  plant,  he  may  find  that  he  is  compelled  to 
close  it,  to  the  ruin  of  the  entire  enterprise.  At  the  Washoe  Works, 
Anaconda,  Mont.,  the  smoke  from  the  furnaces  is  treated  under  the  Cot- 
trell  system  for  the  removal  of  all  dust  and  fume,  so  removing  this  cause 
of  complaint. 

Final  Consideration. — Preliminary  to  building  a  plant  and  operating  a 
works,  an  investigation  is  made  of  the  process,  the  requirements  of  the  plant, 
and  all  limiting  conditions.  It  includes,  besides  the  general  matters 
outlined  above,  the  questions  of  supplies,  markets,  railroad  facilities, 
freight  rates,  sufficient  and  suitable  labor  not  liable  to  strikes,  and  reliable 
civil  conditions  unaffected  by  revolutions  or  oppression  by  the  government 
under  which  the  plant  must  operate. 

Next  comes  the  organization  of  the  operating  company  and  financing 
of  the  enterprise,  or  obtaining  capital  to  build  and  operate  the  plant  until 
it  pays  the  operating  costs. 

Often  the  promoters,  besides  owning  the  mine  for  which  the  reduction 
works  are  built,  have  acquired  the  necessary  real  estate  and  the  rights  that 
go  with  it.  Provisions  should  be  made  for  access  by  railroads,  for  the 
necessary  trackage,  and  for  the  common  roads  to  the  plant.  Not  only 
must  water  and  power  be  provided,  but  right-of-way  for  securing  them. 
If  fluxes  are  needed,  then  the  proper  quarries  or  deposits  must  be  found. 

CONSTRUCTION  OF  PLANT 

Construction. — Before  beginning  construction,  plans  should  be  fully 
worked  out  by  competent  engineers.  Cost  estimates  are  made  in  detail, 
good  materials  are  accumulated,  and  the  labor-force  is  properly  organized. 
In  the  design  of  the  plant,  provision  for  duplicate  parts  is  made,  so  that  in 
case  of  breakdown  no  interruption  of  operation  occurs. 

On  beginning  construction,  the  hydraulic  works,  where  needed,  are  put 
under  skilled  supervision.  This  includes  the  building  of  dams,  reservoirs, 
the  water-power  plant,  and  the  transmission  line.  For  a  long-aistance 
power-transmission  line  there  may  have  to  be  sub-stations  and  a  distrib- 
uting system. 

Money  must  be  provided  for  the  salaries  of  officers  of  the  company 
that  are  to  receive  pay  during  the  period  of  construction,  and  all  money 
expended  must  be  accounted  for,  and  cost-records  kept  by  a  skilled  account- 
ant. The  money  needed  for  legal  expenses,  general  expense,  traveling 
expense,  and  all  expenses  incurred  during  construction  must  be  included. 


CHAPTER  XLV 
ACCESSORY  EQUIPMENT  OF  PLANTS 

Equipment. — This  includes  the  machinery,  furnace-tools,  and  appli- 
ances used  in  operating,  but  excludes  land,  buildings,  and  trackage.  Labor- 
saving  machinery,  when  reliable,  effects  a  saving  in  costs,  but  it  is  remem- 
bered that  this  saving  must  not  sacrifice  the  efficiency  of  operation.  The 
question  "  how  much  "  often  arises,  and  we  may  even  come  to  the  con- 
clusion that  it  is  not  desirable  (considering  the  cost  of  installation)  to  put 
in  the  labor-saving  appliance. 

INTERMITTENT  HANDLING  OF  MATERIALS 

For  handling  on  one  level,  100  tons  or  less  of  material  daily,  especially 
where  the  ore  is  to  be  distributed  to  various 
places,  one  or  two-wheeled  buggies,  or  bar- 
rows, Fig.  295,  on  a  good  floor,  have  been 
found  to  be  economical,  elastic,  and  low  in 
first  cost.     For  small  quantities  the  metal- 
lurgist is  not  led  into  installing  machinery, 
for   he  finds  in  practice  that  it  effects  no 
saving.     For  large  quantities  barrows  or  bug-      FlG  295.-Cha7ging  Buggy, 
gies  may  be  used,  or  hand-propelled  tram- 
cars,  as  in  mining.     For  still  larger  quantities,  power  propelled  cars  are 
used,  that  can  be  handled  also  on  up-grades  and  sent  from  level  to  level. 

The  advantage  of  this  method  of  handling  is  that  loads  can  be  sent  over 
trestles,  above  bins,  can  be  raised  on  elevators,  and  where  they  have  suit- 
able wheels,  can  be  run  over  floors  or  even  upon  the  ground. 

INDUSTRIAL  LOCOMOTIVES 

These  may  be  operated  by  steam,  by  electricity  or  by  compressed  air. 

The  Steam  Locomotive. — Fig.  296  is  a  view  of  one  of  these  suited  to 
the  handling  of  industrial  cars. 

The  Electric  Locomotive. — We  show  in  Fig.  210  an  electric  trolley 
system,  suited  to  the  handling  of  slag  and  matte  on  the  dump,  and  indeed 
for  cars  also.  Fig.  298  is  an  electrically  operated  charge  car,  and  one  is 

551 


552 


ACCESSORY  EQUIPMENT  OF  PLANTS 


FIG.  296. — Steam  Locomotive. 


JO  TON  MINING  TYPE  GASOLINE.  LOCQflOTWE 


FIG.  297. — Gasoline  Locomotive. 


4  7QN  STORAGE.  BATTERY  LOCOMOTIVE. 


FIG.  298.— Electric  Storage 
Battery  Locomotive. 


FIG.  299. — Compressed  Air  Locomotive. 


INDUSTRIAL  CARS  553 

used  at  the  iron  blast-furnace,  Fig.  154,  where  it  is  in  fact  a  moving  weigh- 
scales,  so  that  the  items  of  the  charge  are  weighed  by  it. 

The  Gasoline  Locomotive. — These  are  made  of  sizes  up  to  20  tons  and 
of  any  desired  gauge.  Fig.  297  is  an  example  of  one  of  them. 

The  electric  storage  battery  locomotive,  Fig.  298,  is  a  view  of  such  a 
locomotive.  It  should  be  observed  that  with  it  there  is  no  need  of  a  trolley 
line,  and  that  the  machine  will  traverse  the  entire  yard  trackage. 

Compressed-air  Locomotives. — Air  compressed  at  800  Ib.  per  square 
inch  is  drawn  off  from  an  air-pipe  line  at  a  convenient  point  through  a 
valve  and  hose  into  the  air  tank  of  the  locomotive.  The  supply  will  run 
the  machine  for  several  short  trips  before  it  must  be  replenished.  At  the 
Washoe  plant  of  the  Anaconda  Mining  Company  thousands  of  tons  are 
thus  handled  daily,  and  indeed  the  plant  is  an  admirable  example  of  how  a 
side-hill  site  becomes  effective  when  locomotives  are  used.  Fig.  299 
shows  one. 

INDUSTRIAL  CARS  AND  HOISTS 

For  the  transport  of  material  about  the  works  a  variety  of  cars  are  used 
as  shown  below. 

The  Rocker-side  Dump  Car. — The  Fig.  300  is  a  view  of  one  of  these 


FIG.  300.— Side-dump  Car. 

locked  ready  for  a  load,  also  in  dumping  position.  It  delivers  outside  the 
track  either  to  a  dump  or  into  an  ore  bin. 

These  may  be  provided  with  a  brake,  may  be  built  with  a  scale  for 
weighing  the  contents  of  the  car,  or  may  be  power-driven,  being  provided 
with  an  electric  motor.  Some  again  have  rubber-tired  wheels  for  operation 
about  the  plant  without  the  use  of  tracks. 

Hopper  Cars. — These  (see  Fig.  301)  are  used  for  transferring  calcine 
from  a  roasting  furnace  to  the  hoppers  of  a  reverberatory  furnace  or  of  coal 
to  the  coal  hoppers.  The  bottom  opening  has  a  drop  door  for  the  dis- 
charge of  its  contents. 


554 


ACCESSORY  EQUIPMENT  OF  PLANTS 


Transfer  Cars. — This  is  useful  for  shifting  a  charging  machine  from  one 
side  to  other  of  a  furnace,  or  of  a  charging  vat  (see  Fig.  302).  The 
machine  is  loaded  upon  it  from  one  line  of  track  and  the  transfer  car  takes 
it  over  to  the  other  and  parallel  track. 

Side-discharge  Car. — The  rocker  dump-car  discharges  to  one  side,  the 
side  discharge  car  Fig.  303,  at  both  sides  and  beyond  the  track.  It  is 


FIG.  301.— Hopper  Car. 


FIG.  302.— Transfer  Car. 


FIG.  303.— Side-discharge  Car. 


FIG.  304. — Double-geared 
Platform  Hoist. 


well  suited  for  large  loads  emptying  from  an  overhead  track  into  bins 
below.    The  one  here  shown  is  motor  driven. 

Hoists. — Of  these,  the  commonest  about  reduction  works  is  the  plat- 
form elevator,  Fig.  304,  which  takes  buggies,  wheelbarrows,  or  tram- 
cars  from  floor  to  floor.  It  may  have  a  platform  of  a  size  (6  by  6  ft.)  to 
receive  two  cars  or  wheelbarrows  at  a  time,  and  it  raises  a  one- ton  load 
60  ft.  per  minute.  They  are  often  run  in  balance,  but  it  is  better  to 
have  two  independent  counterweighted  platforms.  Necessarily,  time  is 
lost  in  loading  and  unloading,  so  that  the  estimate  of  the  capacity  is  25 
tons  hourly. 


TRACK  CRANE 


555 


Skip  Car  (Fig.  305). — This  is  used  for  hoisting  material  by  a  steep 
incline-track  for  charging  to  a  furnace.     One  is  shown  as  part  of  the 
equipment  of  an  iron  blast-furnace,  Figs.  153 
and  154.     It  has  a  bale  for  attachment  of  the 
hoisting  rope. 

The  capacity  of  the  skip  is  2  to  5  tons  of  ore 
or  half  the  quantity  of  coke,  and  the  skips  are 
run  in  balance.  It  takes  thirty-four  seconds 
actual  time  for  raising,  dumping,  and  return- 
ing the  skip  to  pit;  but  the  total  time  includ- 
ing the  waits  is  four  minutes,  this  furnishing 
the  supply  to  a  furnace  producing  350  to  500  tons  of  pig  iron  daily  from 
a  total  burden  of  1150  to  1650  tons. 


FIG.  30o. — Skip  Car. 


GRABS  AND  EXCAVATORS 


Grabs. — In  large  establishments  hoisting  rigs  are  used  that  are  pro- 
vided with  large  clam-shell  buckets  or  grabs.  They  take  5  to  10  tons  of 
ore  at  a  time,  and  are  used  for  unloading  vessels,  and  for  transferring  ore 


FIG.  306. — Track-crane  and  Grab-bucket. 


to  stock-piles  for  storage,  or  to  the  furnace  storage-bins,  as  desired.  It  is 
noticed  that  the  movable  frames  or  bridges  are  made  heavy  to  carry  the 
large  loads  safely. 

Telpherage  refers  to  the  transport  by  a  monorail  about  a  plant,  and 
wherever  such  a  rail  can  be  run  there  material  can  be  transported.    We 


556 


ACCESSORY  EQUIPMENT  OF  PLANTS 


give  the  details  of  a  telpher  for  the  transfer  of  matte  pots  from  the  fur- 
naces to  the  yard. 

The  Traveling  Crane. — Fig.  308  gives  in  elevation  a  traveling  crane 
as  commonly  used.     It  is  for  handling  ladles  and  converts  as  described 


FIG.  307.— Telpher. 

under  copper-converting  and  for  Bessemer  and  open-hearth  practice. 
For  handling  materials  inside  a  building  it  is  coming  into  general  use. 
It  is  operated  by  electricity,  and  moves  in  any  direction,  hori- 
zontally or  vertically,  over  the  floor  of  the  building  commanded  by  it, 


FIG.  308. — Traveling  Crane. 


and  it  avoids  obstacles  on  the  floor.  Provided  with  large  mushroom- 
shaped  electro-magnets,  it  is  now  used  to  unload  pig  iron  or  handle 
steel  sheets  weighing  a  ton  or  more,  and  by  using  magnets  no  time  is 
lost  as  in  older  methods  in  passing  chains  around  objects  to  be  lifted. 


BELT  ELEVATORS 


557 


CONTINUOUS  HANDLING  OF  MATERIALS 

Machines  of  this  kind  carry  a  distributed  load,  so  that  the  sub-structure 
upon  which  they  rest  is  light  compared  with  one  upon  which  the  load  is 
concentrated  as  in  a  car.  They  deliver  material  continuously,  and  no  time 


FIG.  309.— Belt-elevators. 

is  lost  in  loading  and  unloading.  Intermittent  conveying,  on  the  contrary, 
if  we  increase  the  load  cf  the  skip  or  bucket,  becomes  slow  and  awkward, 
whereas  in  the  continuous  conveyor  it  is  possible  to  increase  the  capacity 
by  widening  the  conveyor  and  providing  the  correspondingly  increased 
feed. 

We  divide  continuous  machines  into  elevators,  conveyors,  and  con- 
veyor-elevators. 


558 


ACCESSORY  EQUIPMENT  OF  PLANTS 


FIG.  310. — Methods  of  Feeding  Elevators. 


FIG.  311. — Single-strand  End- 
less-chain Elevator. 


FIG.  312.— Double-strand  End- 
less-chain Elevator. 


BELT-CONVEYORS 


559 


Elevators  are  used  for  vertical  or  nearly  vertical  lifting.  The  belt 
elevator,  Fig.  309,  is  of  this  type,  and  consists  of  an  endless  belt  having 
sheet-steel  buckets,  attached  by  flat-headed  elevator-bolts  at  18-in.  inter- 
vals. To  allow  for  the  stretching  of  the  belt,  the  lower  pulley  shaft  is  car- 


FIG.  313. — Screw-conveyor  (quarter  turn). 


FIG.  314. — Discharge  at  Head  of  Conveyor. 

ried  in  take-up  boxes,  by  which  the  shaft  can  be  raised  or  lowered.  The 
lower  pulley  is  enclosed  in  a  boot,  the  ore  delivering  into  the  buckets  at 
the  rising  side  at  the  left.  Ore  not  caught  by  the  buckets  falls  into  the 
boot  and  is  there  scooped  out  by  buckets,  and  is  delivered  to  the  discharge 
spout  by  centrifugal  action  as  the  buckets  pass  over  the  top  pulley. 

Fig.    311    represents    a    single-strand    endless-chain    elevator.     The 


560 


ACCESSORY  EQUIPMENT  OF  PLANTS 


chain  is  carried  by  head  and  foot  sprocket-wheels  with  sprockets  spaced  to 
take  links  of  the  chain.  In  this  case  the  "  take-up  "  of  the  single-strand 
elevator  is  carried  at  the  boot;  for  the  double-strand  one  it  is  at  the  movable 
Upper  shaft  in  the  left.  The  ore  spills  into  a  chute  between  the  two  upper 


FIG.  315. — Movable  Tripper  Discharging.  ^ 

pulleys,  and  in  this  way  the  elevator  can  run  at  the  low  velocity  suited  to 
the  type. 

The  Worm  or  Screw-conveyor. — This  is  convenient  for  delivering 
crushed  ore  or  pulverized  coal  short  distances,  and  it  thoroughly  mixes  the 
ore  conveyed.  Fig.  313  represents  a  screw-conveyor  delivering  ore  from 
the  trough  along  which  it  has  been  conveyed,  into  another  at  right  angles. 
The  ore  drops  from  the  first  to  the  second,  and  is  conveyed  by  the  screw 


BELT-CONVEYORS  561 

in  the  second,  shown  at  the  left.  The  bottom  of  the  trough  is  lined  with 
smooth  sheet-steel  bent  to  conform  to  the  worm  or  screw.  A  screw- 
conveyor  is  shown  in  Fig.  268.  The  disadvantages  of  this  type  of  conveyor 
are,  that  much  power  is  needed,  and  that  the  ore  grinds  on  the  conveyor, 
resulting  in  wear. 

Belt-conveyors  are  used  for  the  horizontal  transfer  of  materials,  and  can 
be  modified  easily  to  carry  up  an  incline.  Of  all  conveyors,  the  belt-con- 
veyor is  most  widely  used.  To  give  it  capacity,  it  is  troughed  by  running 
on  pulleys  that  raise  the  edges  of  the  belt  forming  a  shallow  trough  (see 
Fig.  315.  The  simplest  form  is  an  endless  belt  running  over  end-pulleys, 
the  load  being  fed  at  one  end,  delivering  into  a  chute  or  into  a  bin  at  the 
other.  The  conveyor  carries  a  load  not  only  on  a  level,  but  on  as  steep  as 
24°  incline.  The  capacity  is  large  and  the  conveyors  are  simple  and  dur- 
able. A  12-in.  belt,  traveling  at  the  rate  of  150  to  350  ft.  per  minute, 
delivers  10  to  35  tons  per  hour.  A  24-in.  belt,  traveling  at  the  extreme 


FIG.  316. — Incline  to  Level  with  Movable  Tripper. 

velocity  of  600  ft.  per  minute,  has  a  capacity  of  250  tons  per  hour  of 
crushed  ore,  and  requires  6  H.P.  per  100-ft.  length,  the  power  needed 
varying  with  the  length  of  the  belt.  When  the  ore  ascends  an  incline  we 
add  the  power  for  lifting  the  load.  Fig.  314  shows  a  belt  in  action  deliv- 
ering its  load  over  the  head  pulley  into  an  ore  bin. 

The  Movable  Tripper. — It  is  desired  at  times  to  deliver  the  ore  into  bins 
situated  at  different  points  along  the  belt.  This  is  accomplished  by  using 
the  movable  tripper  shown  in  Fig.  315,  which  also  shows  the  belt  loaded 
with  ore.  To  discharge  the  ore  the  belt  goes  around  the  upper  pulley, 
as  shown,  then  around  a  second  one  just  below,  and  continues  the  course 
to  the  front  end-pulley.  The  ore  shoots  from  the  belt  into  spouts,  that 
deliver  on  either  side  of  the  track  upon  which  the  tripper  moves  into  one  of 
the  bins  below.  The  tripper  can  be  moved  on  its  track  and  set  to  deliver 
to  any  desired  bin. 

At  Fig.  316  we  show  a  typical  installation,  in  which  a  feed  hopper 
at  the  right  delivers  to  an  inclined  belt  that  changes  to  a  level  one,  the  hori- 
zontal part  having  the  movable  tripper.  Beneath  the  belt  is  an  idler  pul- 
ley by  whose  vertical  movement  the  slack  of  the  belt  is  taken  up. 


562 


ACCESSORY  EQUIPMENT  OF  PLANTS 


Endless-chain  Conveyors. — These  are  much  used,  since  they  convey 
ore  not  only  on  a  level,  but  vertically  if  necessary.  Being  entirely  of  metal, 
they  successfully  convey  hot  materials. 

Fig.  317  represents  an  endless-chain  conveyor,  consisting  of  a  series 
of  plates  or  "  nights,"  attached  to  a  double  endless-chain  carried  at  each 
end  by  sprocket-wheels  like  the  double  endless-chain  elevator,  Fig.  312. 
The  ore,  drawn  from  any  desired  storage-bin  as  shown  in  the  figure,  is 
pushed  up  an  incline  by  the  moving  nights  in  a  fixed  steel-lined  trough, 
and  is  taken  by  a  double-strand  endless-chain  elevator  to  a  floor  above.  If 
desired,  slides  may  be  provided  in  the  bottom  of  the  trough.  When  the 
slide  is  opened,  the  ore  drops  into  the  desired  bin  beneath. 

Sometimes,  in  place  of  flights,  a  continuous  series  of  buckets  or  trays  is 


FIG.  317. — Endless-chain  Conveyor. 

used.  These  overlap  so  that  the  ore  cannot  drop  between  them.  They 
operate  upon  the  principle  of  the  Heyl  and  Pattison  pig-casting  machine, 
Fig.  162.  Indeed,  chain  conveyors  lend  themselves  to  a  great  variety  of 
applications,  as  the  examination  of  a  catalogue  of  elevating  and  con- 
veying apparatus  will  show.  The  chief  drawback  to  them  is  that  they  have 
numerous  joints  to  wear,  and  that  the  troughs,  flights,  or  buckets  are  sub- 
jected to  serious  wear.  They  must  run  slower  than  the  belt-con veyoV. 

In  Fig.  74  (elevation)  the  Edwards  roasting-furnace,  we  have  an  example 
of  a  swinging  push-conveyor.  The  flights  are  bladed  and  so  hinged  from  the 
vibrating  carrying-beam  as  to  swing  over  the  ore  in  the  conveying  trough 
on  the  backward  motion,  but  to  push  the  ore  along  when  moving  forward. 
As  is  seen,  the  bottom  of  the  trough  is  provided  with  slides  to  deliver  the 
ore  where  it  is  needed. 


CHAPTER  XLVI 
ORE  STORAGE  AND  SUPPLY 

PROVISION  FOR  SUPPLY 

Ore  may  be  stored  upon  the  ground  or  a  floor,  sometimes  fenced  in  to 
form  a  ground  bin.  Often  for  convenience  in  discharging  into  a  car  or 
conveyor  a  bin  may  be  set  high  enough  to  discharge  into  a  car  or  upon  a 
conveying  belt.  Bins  may  be  large  enough  to  give  a  day's  supply  or  less, 
and  are  then  called  "  feed  bins  ";  or  there  may  be  a  row  or  series  of  them 
for  the  storage  of  various  materials  needed  in  the  milHng  or  smelting  opera- 
tions, and  enough  for  a  number  of  days.  At  the  large  iron  works,  receiving 
ore  from  the  lower  Lake  ports,  it  is  necessary  to  carry  a  winter's  supply 
piled  upon  the  ground.  Such  supply  for  the  furnace  is  picked  up  by  grabs 
and  transferred  to  feed-bins.  At  the  custom  copper  and  silver-lead 
smelting  works  of  the  Western  United  States  a  supply  is  intended  to  last 
from  two  to  six  weeks  and  this  locks  up  much  capital  while  the  ore  is  in 
process  of  treatment.  Even  at  smelting  plants,  treating  their  own  ores,  it 
pays  to  use  from  large  ore  beds  which  have  been  so  stored  as  to  be  quite 
regular  in  composition.  Mills  aim  to  have  in  their  feed-bins  enough  to 
carry  them  on  overnight,  the  coarse  crushing  being  done  on  the  day  shift 
only,  or  in  case  of  a  breakdown,  at  the  supply  end. 

Ore  Bins  or  Pockets. — These  may  be  flat-bottomed  or  inclined-bottom 
bins  and  may  be  made  of  wood  or  of  steel.  When  they  are  flat  the  discharge 
point  is  at  the  bottom  and  side  of  the  bins,  then  when  discharged  about  half 
of  the  contents,  forming  a  natural  slope,  remains.  When  the  bottom  slopes 
three  ways  to  the  side  chute  most  of  the  ore  runs  out.  This  kind  of  bin  is 
shown  in  Figs.  15 A,  40,  45  and  98.  Sometimes  the  bottom  slopes  four 
ways,  forming  an  in  verted  pyramid.  Then,  the  slopes  being  steeper,  the 
discharge  is  better.  At  times  the  bin  may  be  of  steel,  tall  cylinders  on  end 
with  a  side-opening  at  the  bottom,  and  the  ore  forming  its  natural  slope. 

FEEDERS 

Feeders. — There  are  feeders  of  many  types  calculated  to  give  a  speedy 
delivery  as  into  cars  or  upon  a  conveyor;  or  a  regular  or  even  supply  to  a 
furnace,  to  a  crushing  machine  or  to  a  vat:  They  include: 

563 


564 


ORE  STORAGE  AND  SUPPLY 


Ore  Bin  Gates. — The  chute  of  this  gate,  Fig.  318,  is  a  continuation  of 
the  sloping  bottom  of  the  bin.  In  operation  the  attendant  opens  the  gate 
according  to  the  supply  he  needs-  whether  to  quickly  fill  a  car  or  to  supply 

a  crusher.  He  must  watch  for  a  sudden 
rush  of  ore,  or,  when  this  hangs  up,  he 
uses  a  bar  to  loosen  the  ore  and  to 
make  it  run.  Often  there  are  men 
stationed  above  who  poke  down  the  ore 
in  case  the  bin  is  to  be  completely 
emptied. 

Shaking    Screen    and    Feeder. — In 


FIG.  318.— Single  Rack  Gate. 


FIG.  319. — Combination  Shaking-screen 
and  Feeder, 


Fig.  134,  at  a,  is  shown  a  grizzly,  Fig.  54,  to  remove  the  fines  while  the 
lump  ore  is  crushed.  A  better  separation  and  a  regular  feed  is  obtained  by 
substituting  for  this  the  eccentric  driven  screen  feeder,  Fig.  319. 


Ore  Bin 


O    Fines 
B.OTAPY  GRIZZLY  («) 


TRAVELING  FEEDER   (b) 

FIG.  320. — Moving  Feeders. 


Rotary  Feeder. — In  Fig.  320  (a)  is  shown  a  side  elevation  of  a  feeder 
disk  in  which  as  in  the  shaking  screen  there  is  a  separation  into 
fine  and  coarse,  the  latter  passing  on  to  the  coarse  crusher,  while  the 


FEEDERS 


565 


fine  drops  between  the  slowly  revolving  disks.     It  uses  a  minimum  of 
space  horizontally. 

Traveling  Apron  Feeder. — This,  as  shown  in  Fig.  320  (6),  takes  a 
mixed  feed  whose  amount  is  regulated  by  a  slide  at  the  front  of  the 
bin-opening.  It  effects  the  same  separation  of  the  ore  according  to  fine- 
ness. The  fine  ore  is  caught  in  a  by-pass  chute,  which  carries  it  to  clear  the 
lower  chain.  Often  a  tight  apron-feeder  is  used,  delivering  all  to  a  car,  a 
conveyor,  or  to  a  crusher.  Where  the  ore  is  fine,  or  wet  or  sticky,  or  con- 
tains large  lumps,  making  it  liable  to  bridge  or  hang  up,  then  steeper  slopes 
than  the  usual  45°  to  one  side  of  the  bin  are  preferred.  Thus  a  bin  of  wood 
may  have  an  inverted  pyramid  bottom,  or  in  a  steel  bin  a  hopper  of  in- 
verted conical  form.  This  brings  its  discharge  at  a  middle  point  of  the 
ore  column,  and  so  gives  a  surer  run  of  the  ore.  To  take  this  discharge 
the  feeder  is  well  adapted.  The  bottom  opening  can  be  lengthened  out 


FIG.  321. — Reciprocating  Plate-feeder. 


FIG.  322. — Hammer  Feeder. 


liberally  above  the  apron.  It  is  also  well  suited  to  coarsely  crushed  ore, 
especially  for  tall,  narrow  feed-hoppers  that  need  a  well-assured  feed. 
Such  hoppers  used  as  roasters  and  receiving  a  wet  sticky  concentrate  such 
as  is  produced  in  flotation,  are  especially  liable  to  hang  up.  We  may  note 
that  this  type  of  feeder  cuts  down  head  room  to  a  minimum. 

Reciprocating  Plate  Feeder. — This  is  given  in  Fig.  321.  It  receives  a 
quick  reciprocating  motion  from  an  eccentric  and  delivers  through  a 
wide  front  opening,  the  flow  being  regulated  by  swing  hammers,  Fig.  322, 
set  to  allow  the  passage  of  any  large  lump  then  to  fall  back  in  place  again. 
Thus  large  lumps  cannot  jam  the  opening. 

Removal  of  Waste  Wood  and  Tramp-iron  from  the  Feed. — In 
mining  the  ore  chips,  wedges,  and  wood  fragments,  as  well  as  nails, 
chains,  nuts,  and  bolts,  called  tramp  iron,  are  to  be  found  in  the  ore  and 
provision  must  be  made  to  eliminate  these  objects  as  speedily  as  possible, 
especially  the  iron,  which,  if  large  enough,  may  stall  or  break  the  crushing 
machines  that  follow. 


566  ORE  STORAGE  AND  SUPPLY 

In  sampling  ore  the  attendant,  as  already  stated,  regulates  the  flow  of 

ore  from  the  feed  bins.     At  the  same  time  he  removes  the  waste  wood 

and  tramp  iron.     The  former  is  less  dangerous  to  the  coarse  crushing 

machines,  so  that  but  £  or  TV  of  the  whole  is  passed  on  for  finer  crushing. 

Fig.  323  shows  how  the  iron  is  removed.     The  ore  is  carried  from  the 

feed  bins  by  conveying 
belt  to  deliver  to  the 
feed  hopper  of  the  crush- 
ing machine.  The  head 
pulley  of  the  conveyor  is 
magnetized  and  attract- 
ing the  iron  removes  it 
from  the  ore  stream  to 
drop  from  it  as  the  belt 
conveys  it  from  the  mag- 
netic field. 

The  wood  that  passes 
through  the  machines  is 

ground  small  and  at  the 
FIG.  323. — Ding's  Magnetized  Pulley.  ^  n.r        r         a 

first   settling  box   floats 

on  the  water  to  be  removed  by  hand  from  time  to  time. 

PUMPS  AND  ELEVATORS 

For  conveying  concentrate  containing  from  6  to  10  per  cent  moisture, 
belt  and  push  conveyors  are  used,  while  for  elevating  purposes  belt- 
bucket  elevators  are  common.  Elevators  are  a  source  of  trouble,  and  if 
possible  should  be  avoided. 

A  variety  of  different  pumps  is  on  the  market  for  elevating  pulp,  but 
those  that  are  really  good  are  few.  Machines  used  for  this  purpose  are  the 
Frenier  spiral,  centrifugal,  three-throw  plunger,  air-lift,  bucket-elevator, 
and  tailing-wheel. 

The  Frenier  is  satisfactory  for  low  lifts  up  to  8  to  10  ft. 

A  three-throw  plunger  pump  will  give  trouble  with  sandy  material. 
The  valves  cut  out  quickly  and  the  packing  requires  renewing  often.  Air- 
lifts are  simple,  but  require  large  quantities  of  air  for  ordinary  lift£  of  8  to 
10ft. 

The  old  tailing-wheel  gives  the  least  trouble  of  all.  It  is  reliable  and 
the  cost  of  repairs  is  low.  The  one  objection  against  it  is  the  high  first 
cost  to  install. 

For  moving  sand  a  well-designed  centrifugal  pump  with  white  cast- 
iron  liners,  easily  accessible  for  replacing  worn-out  parts,  will  give  good 
satisfaction;  the  Byron-Jackson  pump  is  an  example. 


CENTRIFUGAL  PUMPS 


567 


For  the  movement  of  sand  and  slime  pulps,  water  and  solutions,  both 
low-  and  high-pressure  pumps  are  used.  For  pressure  filters  and  high  lifts 
the  three-throw  plunger  pump  is  much  employed  and  the  centrifugal  pump 
for  large  volume. 

The  Centrifugal  Pump. — Fig.  325  is  a  view  of  this  type  of  pump,  taking 


A   Case 

B    Runner 

C  Suction  Elbow 

D  Suction  Side  Liner 

E  Shaft  Side  Liner 

F  Stuffing  Box 

G  Shaft  Sleeve 

H  Cage  Ring 

J    Gland 

K    Ring  Oiling  Bearing 

L   Oil  Collars  _ 

M  Pulley 

N   Shaft 

O  Braes  Plate 


SECTIONAL   ELEVATION   OF 
1911   STD.   SPLIT  SLIME  PUMP 


FIG.  324. — Centrifugal  Pump.     (Section.) 

its  suction  at  the  front  and  delivering  to  the  down-turned  discharge  pipe 
at  the  right  end.  Having  no  valves  it  works  well  for  pumping  sand  or 
slime  pumps.  Fig.  324  shows  its  internal  construction.  The  rapidly 
revolving  impeller  throws  the  water  to  the  exterior  of  the  casing,  forcing 
it  out  of  the  casing,  when  it  escapes 
to  the  down-turned  discharge  exit. 

Solution  Pumps. — Centrifugal, 
three-throw  plunger,  and  air-lift 
pumps  are  in  common  use  for  ele- 
vating solutions. 

Centrifugals  are  probably  used 
more  than  any  other  for  elevating 
to  heights  of  10  and  up  to  50  ft. 
They  require  considerable  attention 
owing  to  their  high  speed. 

A  three-throw  plunger  will  pump 
solution  to  a  height  of  100  ft.  or  more. 


FIG.  325.— Jackson  Centrifugal  Pump. 
They  require  little  attention,  and 


are  economical  hi  every  way.     I  prefer  them  to  centrifugal  pumps. 


568 


ORE  STORAGE  AND  SUPPLY 


The  air-lift  is  suitable  for  lifts  not  exceeding  10  ft.,  and  is  particularly 
useful  about  a  plant  using  the  counter-current  decantation  process  where 
the  lift  would  not  exceed 
2  to  3i  ft. 

The  Frenier  No.  1  Sand 
Pump. — Fig  .  326  shows 
a  section  of  the  pump.  1«" Service  Cock< 


FIG.  326. — Section  of  Frenier  Pump. 


FIG.  327. — Frenier  Pump  Arrangement.    . 


The  trunk  or  body  of  the  pump,  44  in.  diameter,  mounted  on  a  horizontal 
shaft,  constitutes  a  spiral  rectangular  tube.     There  are  no  valves,  but  the 
sand  and  water  scooped  up  at  each  revolution  of  the  spiral  and  by  the  hy- 
drostatic head  created  by  the  revolution 
flows  to  the  center  of  the  pump,  and  dis- 
charges under  pressure  up  the  discharge 
pipe   at  the   right.     The  spiral  passage 
with  an  opening  2J  by  6  in.  and  at  20 
R.P.M.  will  lift  3000  gal.  per  hour  to  the 
height   of    14  ft.     Due  to  its   simplicity 
and   ease  of  repair  this  pump  is  much 
liked.     Where  it  is  desired  to  increase  the 
height  of  delivery  this  may  be  done  by 
introducing  an  air  jet  into  the  discharge 
pipe  as  shown  in  Fig.  327. 

The  Three-throw  Plunger  Pump. — 
The  plunger  acts  on  the  down  stroke  only 
to  press  out  the  pulp  or  the  solution,  and 
the  grit  cannot  get  past  it  to  cut  the 
cylinder.  Having  three  cylinders  there 
are  three  even  impulses  per  revolution.  An  air  chamber  at  the  discharge 
side  as  seen  in  the  figure  also  tends  to  equalize  the  flow,  Fig.  328. 


FIG.  328. — Triplex  High-pressure 
Pump. 


CHAPTER  XLVII 

COST  OF  PLANT  AND  EQUIPMENT 
COST  OF  PLANT 

Based  upon  figures  of  1913  to  1915,  when  prices  were  comparatively 
stable,  we  give  data  that  may  serve  to  indicate  the  costs  of  metallurgical 
olants.  We  may  safely  calculate,  however,  that  the  costs  in  1920  will  be 
double  these,  but  that,  when  the  present  abnormal  labor  costs  again  return 
to  the  older  figures,  then  plant  costs  will  be  correspondingly  decreased.* 

In  earlier  times,  when  not  so  much  was  done  automatically,  when  wood 
instead  of  steel  buildings  prevailed,  and  where  bedding  was  done  on  floors 
instead  of  in  overhead  bins,  the  cost  would  have  been  half  that  just  given. 
There  are  drawbacks  to  much  permanent  construction  in  which,  where 
changes  are  to  be  made  (and  this  is  often  what  should  be  done),  such 
changes  are  expensive  as  compared  with  those  in  lighter  construction. 

The  first  step  in  such  construction  is  to  obtain  the  services  of  a  com- 
petent constructing  engineer  experienced  in  the  planning  and  building  of 
the  kind  of  works  contemplated.  Such  service  is  particularly  valuable 
in  the  avoiding  of  expensive  alterations,  and  may  amount  to  3  to  5  per 
cent  of  the  total  costs.  It  is  so  much  easier  to  make  changes  in  the  plans 
than  to  later  correct  them  in  the  works  themselves.  The  desire  "  to  make 
the  dirt  fly  "  should  be  overcome. 

Having  matured  the  plans  in  detail  and  made  estimates  of  costs,  based 
upon  the  unit  costs  as  below  given,  one  can  obtain  bids  from  manufac- 
turers of  machinery  and  dealers  in  supplies  and  is  then  in  position  to 
proceed  with  actual  construction. 

To  this  end  unloading  facilities  should  be  provided,  a  good  road  and  if 
possible  the  railroad  tracks  brought  to  the  site  of  the  works.  Ample  room 
should  be  provided  for  lumber  and  for  piling  it  so  as  to  show  just  where 
it  is  to  go.  Also  suitable  storerooms  and  workshops  for  the  use  of 
the  mechanics  are  to  be  built.  Roomy  framing  plots  and  handy  places 

*  It  was  thought  that,  as  after  the  Civil  War  of  1861  to  1865,  prices  would  go 
down  in  the  course  of  two  years,  but  we  must  recollect  that  to-day  European  nations  are 
too  short  of  capital  and  too  exhausted  by  the  recent  strife  to  be  able  to  dump  goods 
upon  our  shores  in  large  quantity,  so  in  the  United  States  labor-demand  bids  fair  to 
keep  up.  When  prices  of  labor  and  supplies  go  down  then  we  may  expect  an  increase  in 
mining. 

569 


570  COSTS  OF  PLANT  AND  EQUIPMENT 

for  the  storage  of  machinery  and  its  protection  against  damage  and  rust  are 
also  to  be  arranged  for.  A  supply  of  water  must  be  available  and  often 
one  may  get  a  supply  of  electricity  for  power  and  lighting  also. 

The  labor  supply  must  be  studied  and  provision  made  for  the  comfort 
and  efficiency  of  the  men.  Without  such  consideration  the  building  of  a 
mill  in  an  out-of-the-way  place  would  prove  disastrous  to  an  enterprise. 

Preliminary  Work. — The  cost  of  all  this  preliminary  work  will  amount 
to  5  to  10  per  cent  of  the  total  and  may  be  estimated  on  the  ground. 

For  concrete  work  a  quarry  may  have  to  be  opened  and  a  bed  of  suit- 
able sand  may  be  available.  Indeed  it  may  be  prudent  to  file  upon  a 
claim  covering  their  location. 

On  page  547  we  have  already  discussed  the  nature  of  the  site  to  be 
chosen  and  the  rights  to  which  the  company  is  entitled  in  filing  upon  ground 
for  a  proposed  mill  or  smelter  site  in  connection  with  a  mine. 

Carpenter  work  with  a  picked-up  local  crew  will  average  $28  to  $31  per 
thousand  for  framing  and  erecting,  $19  per  thousand  board  feet  for  siding 
and  roofing,  $2.50  a  thousand  shingles  for  shingling,  and  $1.25  per  square 
of  100  sq.  ft.  for  putting  on  corrugated  iron.  The  nails  needed  in  erecting 
would  be  18  to  21  Ib.  per  1000  board  feet,  in  putting  on  siding  and  laying 
2-in.  flooring;  while  for  1-in.  flooring  28  to  32  Ib.  per  1000  board  feet  is 
needed. 

Minor  Items  are  Important. — Thus  considerable  lumber  is  needed  for 
forms  and  for  staging.  The  building  should  be  painted,  fire  protection 
and  heating  arranged  for,  office  and  laboratory  equipment  bought. 

Alterations. — Upon  a  completed  mill  it  may  be  necessary  to  make 
alterations  and  this,  as  experience  shows,  may  amount  to  5  to  15  per 
cent  of  the  total  costs. 

Winter  work  in  the  Northern  United  States  or  in  the  mountains  may 
add  as  much  as  33  per  cent  to  the  total  labor  costs  even  in  a  mild  winter, 
and  in  cold  snowy  weather  such  costs  may  rise  to  50  per  cent.  Concrete 
work  often  costs  35  per  cent  more,  as  complete  arrangements  must  be  made 
for  heating  and  protecting  against  the  frost  until  after  the  preliminary  set, 
after  which  freezing  need  not  affect  it. 

Expense  of  Rebuilding  Old  Works. — As  in  a  new  mill,  the  costs  can  be 
rather  accurately  figured,  but  the  amount  of  hardware  and  lumber,  that  can 
be  used  again  is  often  misleading.  The  costs  of  the  carpenter  work  and  of 
the  reassembled  machinery  will  generally  be  twice  that  of  a  new  plant. 

Underestimates. — These  are  due  to  guess-work,  lack  of  good  organiza- 
tion, omissions  and  changes  in  plans,  neglect  of  preliminary  work,  too 
much  reliance  placed  on  general  figures,  and  inefficiency  of  labor  due  to 
unfavorable  conditions.  Also  must  be  mentioned  the  danger  of  strikes, 
bad  weather  delays,  and  failure  of  railroads  or  supply  houses  to  supply 
material  as  needed. 


COSTS  OF  PLANTS 


571 


Machinery  Prices. — A  reputable  machinery  house  will  give  valuable 
information,  and  no  matter  how  confident  the  constructing  engineer  he 
should  give  careful  attention  to  it.  They  are  willing  to  go  into  details 
with  him.  They  do  not  drop  their  responsibility  when  their  machinery  is 
delivered  and  are  always  desirous  of  protecting  themselves  in  this  way. 

Untried  innovations,  especially  by  a  small  plant,  should  be  avoided. 
Let  it  be  tried  out  by  a  larger  operating  company  and,  if  it  is  fully  proved 
there,  it  can  be  put  in. 

If  the  plans  of  the  works  are  carried  out,  a  good  organization  main- 
tained and  efficient  labor  obtained  and  kept,  then  the  figures  for  con- 
struction will  be  found  a  little  higher  than  actual  costs. 

COSTS  OF  METALLURGICAL  PLANTS 

The  following  table  gives  the  daily  capacity  and  the  total  cost  of  a 
variety  of  plants  based  on  figures  of  1913  and  before.  Costs  are  now 
(1920)  easily  double  these: 


Character  of  Plant. 

Capacity  in  Twenty-four  Hours. 

Cost. 

Iron  blast-furnace  

300  tons  pig  iron  

$      650,000 

Acid  Bessemer  with  four  remelting 
cupolas  and  hot  metal  mixer 

2000  tons  of  steel 

900000 

Acid  open-hearth  ;  ten  50-ton  furnaces 

1000  tons  of  steel  

1,500,000 

Basic  open  hearth  ;  ten  50-ton  furnaces 

1000  tons  of  steel    .... 

1,650000 

Copper  smelting  and  converting  
Silver-lead  smelting  

1000  tons  of  ore  smelted  to  100  tons 
of  45  per  cent  matte  and  this 
converted  to  blister  copper  
500  tons  mixed  lead  over  to  base- 
bullion 

1,250,000 
250000 

Refinery  for  base-bullion 

100  tons  base-bullion     refined  by 

the  Parkes  process 

250000 

Refinery  for  dore"  bars  

30,000  oz 

20000 

Refinery  (electrolytic)  

100  tons  copper  from  blister-copper 
to  wire  bars  

500,000 

Zinc  smeltery 

100  tons  of  blende  (not  making  sul- 

Gold mill  (amalgamation)  
Gold  mill  (cyaniding) 

phuric  acid)  
100  tons  of  ore  
100  tons  of  ore 

375,000 
50,000 
100  000 

Copper  blast-furnace  works  

900  tons  (no  roasters) 

560000 

Copper  blast-furnace  works  with  con- 
verter plant            

1200  tons 

984000 

Copper  blast-furnaces  and  converting 
(Washoe  Works,  Anaconda)  

8330  tons. 

10680000 

572 


COSTS  OF  PLANT  AND  EQUIPMENT 


UNIT  CONSTRUCTION  COSTS  IN  1914 

These  are  the  most  useful  to  the  engineer,  since,  having  made  plans  of  a 
plant  or  of  any  proposed  building  he  can  use  these  units  in  making  his 
estimates  of  costs.  These  data,  found  also  in  engineering  hand  books,  are 
carefully  set  forth  for  a  smelting  plant  in  an  elaborate  paper  by  E.  Horton 
Jones,  Trans.  A.  I.  M.  E.,  XLIX,  3.  These  figures,  quite  applicable  to 
the  Clifton,  Ariz.,  district  in  1914  would  need  to  be  doubled  to  conform  to 
our  present  50-cent  dollars,  and  should  be  modified  by  any  accessible 
recent  costs.  At  that  tune  common  labor  cost  $2  and  skilled  labor  $4 
per  day,  as  compared  with  something  like  double  that  now.  It  is  to  be 
noted,  that  the  figures  below  given  for  unit-costs,  are  averages. 

The  works  cost,  completed,  $2,105,020.17.  Out  of  this  has  to  be 
reckoned  $100,649.88  for  engineering  and  $140,277.72  for  indirect  expense, 
including  all  necessary  for  clearing  and  preparing  the  site  and  its  approaches, 
working  equipment,  personal  injuries,  railroad  transportation  to  employees, 
etc.  This  left  $1,864,092.47  for  the  work  that  would  show  upon  the  com- 
pletion of  the  plant.  The  engineering  cost  was  then  5.4  per  cent  of  this, 
indirect  expense  7.53  per  cent,  a  total  of  12.93  per  cent  to  be  added  to  the 
construction  costs. 

The  following  is  a  recapitulation  of  all  costs,  with  its  list  of  buildings, 
all  equipped: 


No. 

Name  of  Account. 

Total. 

No. 

Name  of  Account. 

Total. 

7100. 
7300 

7400 
7700 

Engineering  expense.  .  . 
Yard  tracks  and  indus- 
trial system  
Receiving  bins  
Crushing  plant  

$100,649.88 

156,326.43 
44,185.06 
9,268.62 

8625 
8700 

8714 

Roaster  dust  cham- 
ber flue  
Boiler     and     black- 
smith shop  
Machine    and     car- 

$12,859.10 
21,449.23 

7800 

Sampling  plant 

34  108  74 

penter  shop 

27  356  27 

7900 

Bedding     plant      and 
bunker  bins 

150,939  05 

8800 
8809 

General  office  
Warehouse  ... 

1,394.95 
13,602  71 

8100 

Roasting  plant 

136734  87 

8819 

Laboratory 

6  144  02 

8120 
8300 
8400 
8420 

Roaster  dust  chamber 
Reverberatory  plant..  . 
Converter  plant  
Converter  dust  chamber 

49,664.76 
328,945  .  02 
216,033.37 
27,813.58 

8840 
8900 
8999 
9000 

Sample  room  
Miscellaneous  acc'ts 
Indirect  expense  .... 
Power-plant  

2,826.11 
37,186.48 
140,277.72 
434,70315 

8500 
8600 

Conveying  system  
Chimney 

45,411.15 
45,471  34 

9060 

Oil  supply  sump  and 
pump  house  

40,611  88 

QftlA 

iq  4x0    7/-j 

8620 

Converter  flue  

7,602.88 

Total  cost  

$2,105,020.07 

The  Unit  Cost  for  Concrete  Foundations. — The  average  for  all  founda- 
tion work  would  be  $3.37  for  labor  and  $5.48  for  materials,  a  total  of  $8.85 


UNIT  CONSTRUCTION  COSTS  573 

per  cubic  yard,  in  place.  For  reinforced  foundations  this  is  much  more 
expensive,  the  labor  cost  being  $5.84  and  the  materials  $7.66,  or  a  total 
of  $13.50  per  cubic  yard  in  place. 

Unit  Costs  for  Concrete  Floors. — These  are  laid  like  a  sidewalk  5  to  6  ft. 
square,  4  to  5  in.  thick  and  with  a  finished  top.  They  are  reckoned  at  a 
cost  per  square  foot  as  follows: 

Plain  concrete  floors — $0.08  for  labor  and  $0.13  for  materials  or  a  total 
of  $0.21  per  square  foot.  Reinforced  concrete  floors — $0.18  for  labor  and 
$0.23  for  material,  or  in  all  $0.42  per  square  foot.  These  are  formed,  rein- 
forced, and  finished. 

Unit  Cost  for  Excavation. — This  depends  on  its  nature,  as  below  given: 

For  shallow  excavation,  using  wheelbarrows  and  slips  or  scrapers,  and 
with  a  "  haul,"  or  distance  to  move  the  materials  less  than  100  ft.,  the 
cost  averaged  $0.81  per  cubic  yard  reckoned  in  place.  When  the  haul 
was  greater  than  100  ft.,  needing  carts,  this  cost  rose  to  $0.95.  Where 
the  ground  was  solid,  needing  some  blasting,  even  with  less  than  a  100-ft. 
haul,  the  cost  was  taken  $0.84  to  $0.93.  Like  the  preceding  when  the 
ground  was  hard,  and  the  haul  (using  carts,  etc.)  over  100  ft.  the  cost 
became  $0.89  to  $1.00  per  cubic  yard. 

Averaging  all  the  unit  costs  for  excavation  we  find  it  to  be  $0.79  per 
cubic  yard. 

Unit  Cost  for  Electric  Lighting. — This  is  based  upon  the  cost  for  wiring, 
and  the  material  for  each  drop  or  light  used.  It  is  averaged  at  $4.84  for 
the  labor,  $5.85  for  the  materials,  or  a  total  of  $10.69  per  drop. 

Unit  Costs  for  the  Erection  of  Machinery. — The  total  cost  of  the 
machinery  is  made  up  of  its  cost  f.o.b.  at  the  factory,  plus  freight  to  its 
destination,  plus  the  cost  of  unloading  and  erecting.  This  may  be  com- 
puted at  so  much  per  hundred-weight.  The  manufacturers  will  quote 
weights  and  prices  at  the  factory,  the  freight  rates  may  be  obtained  from 
the  railroad  schedules,  and  the  erection  costs  are  as  here  presented.  It  is 
useful  to  reckon  prices  per  hundred-weight  from  the  machinery  costs  as 
collated,  according  to  the  nature  of  the  machinery.  The  freight  costs 
vary  with  the  classification. 

The  unit  costs,  both  of  the  value  delivered  at  the  plant,  and  for  erec- 
tion, vary  with  their  nature. 

Group  1  refers  to  engine  machinery  that  needs  to  be  placed,  cleaned, 
adjusted,  and  lined  up.  The  cost  of  the  machines  delivered  at  the  plant  is 
$12.67  per  hundred-weight  or  12.67  cents  per  pound.  Add  to  this  $0.92 
for  unloading  and  erecting,  and  we  have  a  total  cost  for  the  machines  in 
place  $13.59  per  hundred-weight. 

Group  2  is  similar  to  Group  1 ,  but  not  so  heavy  and  takes  proportion- 
ately more  labor  to  put  in  working  order.  The  cost  delivered  is  $8.53, 
and  for  erecting  $1.50,  a  total  of  $10.03  per  hundred-weight. 


574  COSTS  OF  PLANT  AND  EQUIPMENT 

Group  3  is  heavy  machinery  needing  little  labor  in  erecting.  The 
machinery  delivered  cost  $1.04,  and  for  erecting  $0.68,  making  a  total  of 
$12.72  per  hundred-weight  installed. 

Group  4.  This  resembles  Group  3,  except  that  it  is  electrical.  Its 
cost  at  the  works  is  $12.67,  for  erection  $1.63,  or  in  all  $14.30  per  hundred- 
weight. 

In  all  these  cases  the  erection  cost  is  made  up  of  labor  and  the  needed 
small  supplies,  as  cotton  waste,  oil,  small  tools,  etc. 

Unit  Cost  of  Masonry. — This  is  given  for  a  retaining  wall  at  $6.19  per 
cubic  yard. 

Unit  Costs  for  Painting. — For  painting  concrete  the  labor  will  average 
$0.08  and  the  paint  $0.12,  or  a  total  of  $0.20  per  square  yard  for  two  coats 
of  paint.  For  painting  iron,  the  corresponding  items  would  be  for  labor 
$0.10  and  for  materials  $0.15,  or  in  all  $0.25  per  square  yard  for  two  coats. 
Woodwork  is  cheaper,  being  but  $0.10  per  yard  in  two  coats,  while  painting 
sash  and  doors  is  expensive,  being  $0.96  per  sash,  one  door  being  reckoned 
as  two  sashes,  and  all  being  three-coat  work. 

Unit  Costs  of  Roofing. — When  the  roofing  consists  of  1-in.  sheathing 
covered  with  asbestos,  but  not  painted,  the  cost  was  per  square  of  100  sq.  ft., 
for  labor  $4.06,  and  for  materials  $12.40,  or  in  all  $16.46  laid.  Much  of 
the  roofing  was  of  2-in.  stuff  with  asbestos  covering,  taking  a  total  cost 
of  $26.08  per  square.  The  costs  vary  greatly  according  to  the  kind  of 
roofing  needed. 

Unit  Costs  of  Shafting,  Pulleys  and  Belting. — The  basis  is  per  linear  foot 
of  shafting  equipped  with  its  average  of  pulleys  and  belting.  According 
to  the  building  in  which  it  is  used  it  varies  from  $22.76  in  the  sampling 
plant,  to  as  little  as  $7.98  per  foot  in  the  blacksmith  shop. 

Unit  Costs  for  Structural  Steel. — This  has  reference  to  the  steel  used 
in  the  construction  of  trestles,  buildings,  etc.  Like  machinery,  it  includes 
the  cost  laid  down  at  the  works,  plus  the  cost  of  erection.  It  is  a  pretty 
uniform  figure  and  will  average  $87.13  per  ton  or  4.356  cents  per  pound. 

Unit  Costs  for  Ventilators,  Windows  and  Doors,  Woodwork  and  Wooden 
Floors. — Ventilators  cost  on  an  average  $95.67  each,  erected;  windows 
and  doors  were  reckoned  at  $0.81  per  square  foot.  Woodwork  cost  $52.55 
per  1000  sq.  ft.  board  measure,  while  the  wood  floors  cost  $0.21  per  square 
foot  erected. 

Unit  Costs  for  Labor. — Upon  these,  as  shown  above,  other  costs  depend 
and  it  might  be  well  for  a  rough  approximation  to  vary  the  costs  here 
shown  on  the  basis  of  the  labor-cost.  Thus,  with  the  cost  of  labor  doubled, 
we  may  expect  that  supplies  have  correspondingly  increased,  and  so  that 
the  above  estimates  are  to  be  doubled.  The  proper  way  is,  however,  to 
readjust  the  labor  costs  as  given,  by  the  new  figures,  for  labor,  and  ascer- 
tain freshly  the  cost  for  supplies  and  equipment. 


COMPOSITE  COSTS  575 

Wage  Scale.— This  was  in  September,  1913,  at  Clifton,  Ariz.,  for 
common  labor  $2,  for  skilled  labor  $4  per  day  of  eight  hours.  We  may 
note  that  of  late,  due  to  increasing  wages,  and  the  scarcity  of  labor,  men 
have  become  less  efficient. 

COMPOSITE    COSTS 

These  are  convenient  figures  for  arriving  at  an  approximate  idea  of 
the  cost  of  a  building,  either  empty  or  equipped,  based  upon  its  area  or 
cubic  contents  or  of  its  machinery  and  equipment,  according  to  its  capacity. 

The  cost  of  buildings  varies  from  as  little  as  $1.51  for  the  roasting  plant, 
to  as  much  as  $3.62  for  the  crushing  plant  per  square  foot  of  floor  area. 
In  re  erence  to  cubic  contents  the  cost  varies  from  $0.11  to  $0.22  per  cubic 
foot  for  the  respective  buildings  just  quoted. 

When  it  comes  to  the  cost  of  these  buildings  with  all  their  machinery 
or  equipment  in  place,  the  figures  are  increased  according  to  the  cost  of 
that  equipment.  Thus  the  roasting  plant  comes  to  $4.76  and  the  crushing 
plant  to  $5.62  per  square  foot  of  floor  area,  while,  for  cubic  contents,  the 
former  is  reckoned  at  $0.33  and  the  latter  $0.34  per  cubic  foot,  showing 
how  expensive  relatively  is  the  equipment  of  the  roaster-plant. 

The  feed-bins  and  the  bedding-floors  (using  the  Messiter  System)  cost 
$0.66  per  cubic  foot  of  capacity,  the  heavy  receiving-bins  come  to  $3.34  for 
the  same  unit. 

Conveyors  will  cost  from  $19.02  to  $32.58  per  ton  of  hourly  capacity, 
this  including  the  cost  of  the  steel  supporting  structure.  If  we  were 
reckoning  a  conveyor  according  to  its  cost  per  linear  foot,  erected,  this 
would  be  $34.47  per  foot. 

Dust-chambers  are  $0.30  on  an  average  per  cubic  foot,  the  flues  $0.45. 

A  useful  figure  is,  that  the  cost  for  power-house  installation  (including 
the  boilers),  would  be  $55.32  per  indicated  horsepower,  while,  if  the  boiler 
plant  is  not  included,  this  drops  to  $37.40  for  the  same  unit. 

Again,  the  cost  per  ton  of  output  in  twenty-four  hours  for  the  rever- 
beratories  is  $43.47  per  ton. 

The  cost  per  roaster,  complete  with  its  installation,  is  $17,091.86,  and, 
with  its  building  and  flues  and  dust-chambers,  this  figure  rises  to  $24,097.34. 

The  railroad  trackage  cost  $4.64  per  running  foot,  being  $2.71  for  labor 
and  $1.94  for  materials. 

Raw-material  Prices. — In  Mr.  Jones'  paper  a  table  of  such  supplies 
is  given  f .o.b.  Clifton,  during  1913,  and  to  this  the  student  is  referred. 

The  author  is  of  the  opinion  that  a  careful  study  of  this  paper,  including 
its  illustrated  details  and  descriptions,  would  constitute  a  great  aid  in  an 
engineering  education. 


PART  IX 
THE  BUSINESS  OF  METALLURGY 


CHAPTER   XLVIII 
THE  GENERAL  ECONOMIC  SITUATION 

DISTRIBUTION  OF  WEALTH 

So  far  as  the  United  States,  in  its  internal  economic  condition  is  con- 
cerned, we  may  say: 

The  fixed  wealth  of  the  United  States  in  1916  was  about  $260,000,000,000, 
whereof  about  $30,000,000,000  was  in  stocks  of  goods  and  all  the  rest  in 
real  estate,  railways,  etc.  The  population  of  the  country  was  about  102,- 
500,000  souls,  of  whom  about  41,000,000,  men  and  women,  were  workers, 
about  14,000,000  of  them  being  farmers.  The  total  national  produce 
was  about  $1,200,000,000  tons  of  goods,  worth  about  $45,000,000,000 
to  $50,000,000,000.  Out  of  that  produce  a  group  aggregating  a  little 
more  than  400,000,  who  received  incomes  in  excess  of  $3000  and  paid 
income  taxes,  got  about  $7,900,000,000.  Less  than  one-half  of  that  was 
derived  from  investments  and  more  than  one-half  came  from  the  per- 
sonal efforts  of  this  class.  Persons  enjoying  incomes  of  less  than  $3000 
received  about  44  per  cent  of  the  dividends  paid  by  corporations  and  a 
much  larger  proportion,  perhaps  75  per  cent  of  the  government,  state, 
municipal  and  corporate  interest  payments.  There  remained  from  $23,000- 
000,000  to  $28,000,000,000  to  be  divided  among  27,000,000  non-agricul- 
tural workers,  who  received  an  average  of  somewhere  betweeen  $855  and 
$1040  each.  Among  the  great  classes  of  workers  there  is  a  wide  difference 
in  earnings.  The  farm  hand  in  1916  averaged  about  $400,  the  factory 
worker  $675,  the  steam  railway  man  $886,  and  the  metal  miner  $1250. 
Some  classes  probably  averaged  higher  wages  than  the  metal  miner. 

A  satisfactory  economic  system  can  be  based  only  on  natural  human 
impulses,  and  of  these  the  most  fundamental  is  self-interest.  Increased 
production  is  at  the  present  moment  the  most  pressing  national  need,. but 
it  will  become  effective  only  when  for  every  man  increased  production 
becomes  the  talisman  by  which  his  paper  wages  can  be  turned  to  gold. 

ECONOMICS  OF  ENGINEERING 

The  mining  engineer,  entering  upon  the  practice  of  his  profession,  may 
confine  himself  to  the  technique  of  mine-operating,  while  the  ore,  delivered 
from  underground,  is  then  taken  in  charge  by  the  metallurgical  engineer, 

579 


580  THE  GENERAL  ECONOMIC  SITUATION 

whose  business  it  is  to  win  the  metals  from  it.  The  young  engineer,  enter- 
ing metallurgical  practice,  takes  subordinate  work,  such  as  drafting  or 
assaying  or  testing,  which  gives  him  thorough  knowledge  of  certain 
branches  of  the  work  that  he  is  to  take  up  later  in  operating.  On  the  other 
hand,  the  duties  of  metallurgical  practice  may  be  early  assigned  to  either 
the  draftsman  or  assayer,  so  that  they  may  be  compelled  to  think  along  the 
lines  of  actual  practice.  We  find,  as  a  matter  of  fact,  that  men  actively 
operating,  are  thinking  much  of  those  duties,  are  studying  and  discussing 
them,  often  relegating  to  the  background  the  economic  considerations 
later  liable  to  come  up;  hence  the  discussion  given  on  the  following  pages 
to  these  aspects  of  metallurgical  engineering. 

The  Economic  Situation  in  the  United  States  as  Related  to  the  Pro- 
duction of  Metals. — The  prosperity  of  custom  works  reflects  that  of  mining 
and  profits  in  them  fall  away  in  dull  times,  since  their  charges  must  be 
reasonable  in  order  to  get  the  ores.  With  the  mines  plants  it  is  different 
and  their  prosperity  is  tied  up  with  that  of  the  mine.  Milling  or  smelting 
the  ore  is  but  one  item  in  mines  operation. 

The  value  of  a  metal  is  fixed  by  the  cost  of  production  at  the  "  marginal 
class  of  mines,"  that  is  those  mines  that  just  pay  their  way.  If  the  price 
of  the  metal  goes  up,  leaner  mines  may  become  marginal,  while  the  first 
ones  cited  come  into  the  profitable  class.  If  the  price  of  the  metal  drops 
the  marginal  mine  must  close  down. 

THE  LABOR  SITUATION 

As  the  country  has  opened  up  so  has  the  mining  industry,  and  the 
demand  for  labor  has  been  met  in  part  by  European  immigration.  This 
has  of  late  practically  ceased  so  that  workingmen  are  acutely  needed; 
and  wages  have  doubled.  Organized  labor  has  taken  advantage  of  this  to 
make  demands  under  pretense  of  needing  a  "  living  wage,"  that  is  money 
enough  to  meet  the  necessities,  but  also  many  of  the  luxuries  of  life. 
Often  the  strikers  have  not  cared  whether  they  worked  or  not;  a  holiday 
would  well  suit  them.  As  a  result,  capital,  whose  rewards  depend  upon 
uninterrupted  operation,  has  lost  seriously,  and  the  marginal  mines  are  hav- 
ing to  shut  down,  thus  again  cutting  down  the  supply  to  custom  plants. 

Unions  and  Non-union  Labor. — Two  methods  are  in  use  in  industrial 
plants,  viz.,  the  open  shop  and  the  unionized  shop.  In  the  case  of  the 
open  shop  the  method  of  individual  bargaining  prevails;  in  the  second, 
that  of  collective  bargaining,  that  is,  the  agreement  for  wages,  hours,  and 
treatment,  are  made  by  the  officers,  or  by  a  committee  on  behalf  of  the 
men,  who  belong  to  a  labor  union.  In  the  second  method  it  is  expected 
that  none  but  union  men  are  to  be  employed. 

The  union  shop  with  its  working  force  is  controlled  by  the  labor  union. 


THE  LABOR  SITUATION  581 

Workingmen  who  treat  individually,  lacking  the  backing  of  a  union,  may 
be  taken  advantage  of  by  the  employer,  who  offers  them  a  low  wage  or 
treats  them  in  an  arbitrary  way,  and  such  men  wish  the  support  of  the 
union.  The  union  declares  the  equality  of  all  their  men,  says  that  the  fast 
workman  shall  do  no  more  than  the  slow  one,  because  the  fast  man  com- 
pels the  slower  one  to  work  to  exhaustion;  that,  if  output  is  increased,  then 
demand  will  cease  and  the  workman  will  be  out  of  a  job. 

The  worst  feature  of  a  union  is,  however,  its  tyrannical  power,  that  it 
makes  its  demands  on  the  penalty  of  a  strike  to  enforce  them,  that 
the  power  to  call  a  strike  is  entrusted  to  intermediary  officers,  when 
it  is  notorious  that  such  positions  have  been  sometimes  gained  by  cajolery, 
bribery,  and  the  methods  of  ward  politicians.  In  such  cases  strikes  have 
been  called  when  but  a  small  fraction  of  the  working  force  has  desired  them. 
In  a  unionized  works  men  who  are  employed  in  special  work  may  be  unrep- 
resented in  committee.  There  may  be  a  dozen  or  more  of  such  specialized 
positions  enjoying  compensation  dependent  on  their  skill.  There  will 
always  be  an  incentive  on  the  part  of  the  committee  man  to  favor  his  own 
job,  or  his  own  friends,  and  on  the  other  hand  the  works  manager  may  be 
only  too  willing  to  back  the  committee  man  if  he  sees  it  is  to  his  advantage. 
Such  methods  produce  discontent,  and  eventually  a  strike.  Even  men  who 
receive  the  highest  pay  may  be  so  affected.  It  is  suggested  that  these 
highly  paid  men  might  be  paid  even  more  in  order  to  keep  them  quiet, 
but  there  will  come  a  time  when  this  is  more  than  the  union  as  a  whole 
will  stand,  since  the  action  of  the  committee  is  not  final.  The  signature 
of  the  company  bears  with  it  responsibility,  but  the  signatures  of  the  com- 
mittee do  not.  The  union  is  not  incorporated;  it  has  no  tangible 
assets,  it  is  irresponsible.  It  cannot  bind  an  individual  to  work,  and  if 
there  is  a  good  demand  for  labor  he  may  seek  it  elsewhere.  If  indeed  the 
union  wishes  to  aid  some  other  it  may  go  out  on  a  sympathetic  strike,  due 
to  no  fault  of  the  responsible  company. 

Where  large  bodies  of  skilled  men  of  one  trade  join  in  a  union,  that  is 
different,  but  for  a  works  having  varying  pay  according  to  the  skill  of  the 
workman  the  union  is  bound  to  be  inequitable.  The  total  unionized  labor 
in  the  United  States  is  3,000,000  out  of  35,000,000  people  engaged  in  gainful 
occupations;  it  constitutes  a  body  one-twelfth  of  all  the  workers,  holding 
up  production  if  striking,  and  raising  the  cost  of  living  to  all  the  rest, 
and  especially  the  many  income  receivers,  who,  due  to  their  thrift,  have 
invested  in  these  very  companies.  Thus  a  man  who  has  thought  to  pro- 
vide for  his  retirement  in  old  age,  or  to  insure  on  behalf  of  his  loved  ones, 
is  compelled  to  labor  on,  or  to  live  a  constricted  existence. 

The  labor  union  also  makes  the  tyrannical  demand  that  non-union 
men  shall  not  work  with  union  men,  that  the  shop  shall  be  a  "  closed  " 
one,  even  that  goods  made  by  non-union  men  shall  not  be  delivered  to  a 


582  THE  GENERAL  ECONOMIC  SITUATION 

union  shop.  Many  modern  strikes  are  based  on  these  ideas,  and  the 
strikers  are  prepared  to  carry  them  out  with  picketing  or  even  with  deadly 
violence. 

The  union  disapproves  of  labor-saving  machines,  unless  the  profits 
arising  from  their  use  are  distributed  among  all  the  men,  its  reason  being 
that  it  fears  production  will  outrun  demand  and  so  the  workmen  will  have 
nothing  to  do.  A  sufficient  reason  why  such  machines  are  not  contrary 
to  the  interest  of  the  workmen  lies  in  that  fact  that  formerly,  when  such 
work  was  done  by  hand,  a  skilled  man  had  to  be  physically  superior  and 
by  middle  life,  despite  his  skilled  knowledge,  had  to  take  inferior  work. 
With  the  introduction  of  the  machine  he  could  retain  his  employment, 
indeed  earn  more  than  when  he  laboriously  worked.  It  was  of  mutual 
advantage  to  retain  the  services  of  such  an  experienced  and  trusted  man. 

ARBITRATION 

This  implies  the  settlement  of  disputes  between  employer  and  employee. 
By  enactment  of  law  it  may  be  compulsory  or  advisory.  By  private  agree- 
ment an  arbitration  committee  depends  on  its  power  of  persuasion,  or  the 
willingness  of  both  sides  to  submit.  In  this,  public  opinion  may  come  in 
as  a  factor,  especially  if  its  interests  are  directly  involved.  In  so-called 
compulsory  arbitration  one  can  hardly  see  how  a  works  can  be  compelled 
to  operate  at  a  loss  except  by  confiscation,  nor  how  a  man,  or  even  many  of 
them,  can  be  compelled  to  work  if  they  do  not  wish  to;  they  are  pecuniarily 
irresponsible.  To  be  sure,  as  has  often  been  done,  when  the  employer 
has  submitted,  it  has  been  possible  for  him  to  raise  his  prices  to  meet  the 
increased  labor-cost,  and  so  pass  them  on  to  the  ultimate  consumer.  Arbi- 
trating committees  may  be  appointed  to  adjust  grievances,  but  a  better 
way  is  that  each  man  who  feels  he  has  been  ill-treated  should  have  access 
to  the  works-manager  or  superintendent,  the  ultimate  judge.  In  this 
way  favoritism  or  nepotism  may  soon  become  unknown,  and  injustice 
checked  in  its  beginning.  Investigation  shows  in  general  that  the  man's 
statements  are  correct,  and  a  fair  and  equitable  arrangement  should  be 
made.  In  these  cases  it  is  well  that  the  matter  be  discussed  in  private  and 
that,  for  psychological  reasons,  the  complainant  in  discussing  it  should 
sit  down. 

Where  work  is  abundant  and  jobs  waiting,  men  become  arbitrary, 
notional  and  unreasoning,  since  they  feel  they  can  compel.  When  times 
are  dull  and  work  hard  to  find,  they  will  take  what  they  can  get;  the 
employer  may  then  become  arbitrary.  We  give  herewith  the  periods  of 
panics  and  their  causes  which  can  be  used  in  guiding  future  action  on  the 
part,  at  least,  of  the  employer. 


ASSOCIATION  OF  EMPLOYERS  583 

FINANCIAL  CRISES  IN  THE  UNITED  STATES 

First 1819 

Second 1837 

Third 1857 

Fourth 1873 

Fifth 1885 

Sixth 1893 

Seventh ...... 1907 

Eighth — severe  decline  in  business 1913 

Ninth  ? 1927  or  1926 

ASSOCIATION   OF  EMPLOYERS 

We  have  mentioned  the  shortcomings  of  the  workingman;  the  question 
arises:  What  is  the  employer  to  do?  His  best  plan  is  to  imitate  the 
methods  of  the  union.  Association:  This  is  all-important.  Thus  the 
mining  industry,  the  smelting  industry,  the  milling  industry  should  unite  to: 

(1)  Form  a  union  to  which  all  should  liberally  contribute,  probably 
small  and  large  alike. 

(2)  Appoint  an  executive  committee  to  establish  a  propaganda  based 
upon  exact  and  statistical  information.     These  data  should  be  accessible 
to  writers  who  can  make  good  use  of  them  and  these  writers  should  be 
properly  compensated.     There  should  be  statisticians  who  have  power  to 
enter  into  the  matters  of  costs  and  prices,  also  men  skilled  in  deducing  con- 
clusions from  them.     Income-receiving  men  of  limited  means  should  be 
hired  to  preach  this  propaganda,  and  carefully  instructed.     Where  the 
situation  is  fitting  the  officials  of  the  companies  should  speak  at  meetings 
and  elsewhere.     The  association  should  use  the  lockout  judiciously  and 
should  carefully  prepare  for  the  shut-down. 

Finally,  even  as  the  union  man  will  strike  and  starve  to  gam  his 
ends,  so  must  the  company  do,  confident  that  by  lockouts  a  permanent 
improvement  can  be  made.  Where  one  or  two  out  of  many  shut  down, 
it  is  but  fair  they  should  draw  compensation  from  the  general  fund. 


CHAPTER  XLIX 
ORGANIZATION    AND    OPERATING 

ORGANIZATION  OF  A  METALLURGICAL  COMPANY 

Metallurgical  operations  on  a  commercial  scale  require,  generally, 
the  organization  of  a  company,  or  if  the  company  is  already  organized, 
the  establishment  of  a  department  to  provide  the  additional  function. 

Where  a  metallurgical  company  is  to  be  organized,  the  promoters  or 
organizers  obtain  a  charter,  or  articles  of  incorporation,  from  the  State  in 
which  they  desire  to  incorporate.  They  next  hold  a  meeting  at  which  they 
receive  the  property  that  is  to  be  taken  over  by  the  company,  adopt  a  set 
of  by-laws  for  the  guidance  of  the  company,  and  elect  the  directors  that  are 
to  manage  the  affairs.  The  directors  proceed  to  the  election  of  the  cor- 
porate officers  of  the  company  from  their  number.  The  officers  of  a  small 
company  are  the  president,  the  vice-president,  the  secretary  and  the 
treasurer.  The  directors  may  appoint  from  their  number  a  managing 
director,  or  they  appoint  a  manager  from  the  outside  to  have  charge  of  the 
affairs  of  the  company. 

THE  ADMINISTRATIVE  DEPARTMENT 

In  outlining  the  organization  of  a  company  undertaking  metallurgical 
works,  the  manager  should  be  guided  by  the  following  rules : 

He  should  see  that  a  supreme  authority  is  provided  over  all  action  to 
be  taken,  and  should  carefully  and  fully  outline  the  authority  and  respon- 
sibility of  each  position,  making  the  duties  of  each  conform  to  the  capa- 
bility of  the  party  holding  it.  To  do  this,  he  must  avoid  making  any 
person  subordinate  to  two  or  more,  should  place  the  authority  and  respon- 
sibility together;  should  distribute  the  work  and  the  duties  not  Ho  over- 
burden nor  to  underload;  and  should  arrange  the  positions  so  that  pro- 
motion can  come  from  them.  While  the  manager  gives  his  chief  attention 
to  the  commercial  or  business  affairs  of  the  company,  he  generally  appoints 
a  superintendent  to  attend  to  the  technical  affairs  of  the  plant. 

The  industrial  organization,  under  charge  of  a  manager,  would  include 
(1)  The  operating  department;  (2)  the  accounting  department,  and  (3) 
the  purchasing  and  selling  and  supply  department. 

584 


OPERATING  DEPARTMENT  585 

(1)  The  operating  department  has  to  do  with  all  that  pertains  to  the 
reduction  or  manufacture  of  the  ore  into  metal  (the  winning  of  the  metal 
from  ore)  or  to  refining  metals  to  bring  them  into  marketable  form,  and 
has  control  of  the  operating  forces,  consisting  of  the  foremen  (and  men 
under  them),  the  repair  force  (consisting  of  mechanics  and  their  helpers, 
who  keep  the  plant  in  repair  and  put  in  the  needed  improvements),  and  the 
laboratory  or  assay-office  force. 

(2)  The  accounting  department  attends  to  the  accounting,  pay-roll, 
cost-keeping,  and  the  distribution  of  costs. 

(3)  The  purchasing,  selling  and  supply  department  attends  to  the  pur- 
chase of  ore,  fuel,  fluxes,  and  the  chemical  and  other  supplies.     It  sells 
the  products  of  the  works.     By-products,  in  process  of  farther  treatment, 
are  not  included. 

THE  OPERATING  DEPARTMENT 

We  discuss  the  qualifications  of  those  in  charge,  the  management  of  the 
working  force,  their  welfare,  efficiency,  and  their  payment. 

Duties  of  the  Superintendent. — The  superintendent  not  only  must  be 
informed  as  to  the  actual  technical  operations,  but  he  must  know  how  to 
organize  his  force.  He  should  be  able  to  handle  men  effectively  through 
tact,  discretion,  and  firmness.  He  should  be  strict  but  just;  able  to 
encourage  as  well  as  to  drive.  He  is  often  the  metallurgist  and  construct- 
ing engineer  as  well  as  the  superintendent;  and  has  the  direct  manage- 
ment, with  the  aid  of  his  assistants  and  foremen,  of  the  furnaces  and  the 
metallurgical  machinery. 

When  things  go  wrong  he  may  be  called  on  at  any  hour  to  correct  them; 
if  a  furnace  is  in  bad  condition,  or  a  machine  out  of  order,  he  is  responsible. 
When  all  is-  going  smoothly  his  duties  may  be  light,  but  when  troubles 
come,  or  the  company  is  losing  money,  his  work  is  hard.  If  he  fails  in 
adjusting  difficulties,  no  excuse  is  accepted;  he  must  succeed  or  resign. 
Much  of  his  success  depends  on  his  subordinates,  and  first  in  importance 
among  them  the  foremen. 

General  orders,  applying  to  different  departments,  should  be  issued  in 
multiple  so  that  each  foreman  affected  shall  have  a  copy,  thus  avoiding 
delay  or  misunderstanding. 

The  work  of  supervision  and  control,  that  is,  for  the  superintendent,  his 
assistants,  the  foremen,  the  testing  and  laboratory  force,  and  for  the  office 
men  is  paid  monthly. 

The  men  chosen  to  take  charge  of  the  different  departments  should  be 
chosen  according  to  their  qualifications  for  that  particular  work,  essential 
qualities  being  intelligence  and  reliability.  Lower-grade  labor  is  used 
for  plain,  hard,  routine  work,  the  better  labor  where  judgment  is  necessary. 
Common  or  unskilled  labor  is  liable  to  make  blunders;  however,  when 


586  ORGANIZATION  AND  OPERATING 

trained,  if  faithful,  it  becomes  reliable.  The  saving  made  by  the  employ- 
ment of  cheap  men  for  operating  costly  machines  is  offset  by  the  loss  of 
time,  or  by  actual  disaster. 

Duties  of  Foreman. — A  foreman  is  often  one  who  has  been  advanced 
from  a  lower  position  in  the  same  works,  or  he  may  have  been  selected 
from  another  establishment.  With  a  view  to  efficiency  it  is  well  in  select- 
ing a  foreman  to  find  out  from  former  employers  how  he  gets  on  with 
his  men,  whether  he  finds  fault  with  them,  whether  he  has  been  threat- 
ened by  them.  Though  he  may  lead  off  or  show  them  how  things  are  best 
performed,  in  general,  he  has  enough  to  do  in  seeing  that  the  work  is  well 
planned,  and  that  the  working  force  are  busy.  He  has  not  only  to  note 
the  execution  of  the  work,  but  to  plan  ahead,  to  be  sure  that  everything  is 
provided  for,  and  is  at  hand  when  needed.  In  much  of  such  work  he 
need  not  drive  as  in  routine  work. 

Testing  or  Research  Work. — This  department  needs  a  head,  not  only 
capable  of  carefully  making  tests,  but  also  of  drawing  useful  deductions 
from  them.  Thus,  a  test  may  consist  in  determining  which  of  two  or  more 
methods  will  be  the  most  effective  or  economical.  For  example,  one  may 
wish  to  learn  what  coke  (taking  account  of  its  price)  would  cost  the  least 
per  ton  of  charge,  or  which  of  several  methods  of  admitting  air  to  one  or 
other  of  the  hearths  of  a  MacDougall  roaster  will  give  the  best  roast. 

The  chemist  and  assayer  are  called  on  for  the  results  of  the  analysis 
of  by-products  of  the  works,  of  the  ores  purchased,  and  of  the  products  sold. 
He  must  produce  with  promptness  results  that  control  operations.  In  the 
case  of  ores  bought,  and  of  products  sold,  accuracy  is  fundamental.  Cer- 
tain supplies  like  oils  and  chemicals  may  have  to  be  analyzed  by  him. 

Repair  Force. — The  repair  force,  consisting  of  the  master-mechanic 
and  skilled  men  under  him,  not  only  have  to  make  repairs,  but  have  new 
construction  to  attend  to,  generally  under  supervision  of  the  superintendent, 
who  may,  where  the  works  require,  employ  a  constructing  engineer  and 
draftsmen.  It  is  a  good  rule,  in  case  of  a  breakdown,  or  other  similar 
emergency,  that  this  work  have  precedence,  and  other  work  be  dropped  to 
expedite  it.  A  plant  might  easily  be  losing  a  dollar  per  minute  during  such 
a  period. 

Workmen. — The  laborers  at  a  smelting  plant  are  largely  unskilled. 
These  are  called  outside  men  or  "  roustabouts."  They  do  the  wbrk  need- 
ing pick  and  shovel,  such  as  unloading  cars,  handling  the  products  of 
the  works,  carrying  materials,  etc.  Skilled  laborers,  or  men  working  in 
shifts  of  eight  hours  daily,  and  called  inside  men,  receive  higher  pay  per 
day  than  common  laborers,  the  pay  varying  acording  to  the  particular 
position  held.  These  men  are  responsible  for  the  successful  performance  of 
the  duties  assigned  to  them. 


RULES  OF  WORK  587 

RULES  OF  WORK 

For  keeping  discipline,  and  to  prevent  slackness,  certain  rules,  the 
result  of  long  experience,  have  been  laid  down.  These  are: 

The  men  must  be  promptly  at  work  and  must  work  full  time. 

Inside  men  must  be  on  hand  the  entire  time  of  their  shift,  and  must 
eat  their  luncheon  as  they  can  spare  the  time,  while  not  neglecting  their 
duties.  Charge-wheelers  must  keep  up  the  supply,  but  may  rest  at 
intervals  as  they  wish.  They  are  not  called  on  to  do  other  work,  except 
sweeping  up  their  own  places  before  going  off  shift.  The  inside  man  can 
leave  when  relieved  by  his  partner,  but  must  wait  for  the  partner  until 
relieved.  If  the  latter  fails  to  appear  the  foreman  provides  another  man, 
who  then  holds  the  place,  the  absent  man  losing  it,  unless  he  has  a  good 
excuse,  or  if  sick,  he  is  expected  to  notify  the  foreman  who  provides  a  man 
for  the  place.  When  the  absent  man  desires  to  return  to  work,  he  must 
notify  the  foreman  one  shift  in  advance,  so  that  the  substitute  is  not  put 
out  of  a  shift  for  which  he  has  come  prepared. 

When  men  are  sick  on  shift,  if  not  too  seriously,  they  should  be  held, 
if  possible  to  the  end  of  the  shift.  It  is  impressed  on  them  that  it  is  detri- 
mental to  the  work  for  them  to  leave,  and  that  it  is  difficult  at  such 
short  notice  to  get  a  substitute. 

Men  must  obey  orders,  and  disobedience  can  only  be  followed  by  dis- 
charge, otherwise  discipline  is  weakened. 

Let  the  foreman  be  strict  but  just.  It  helps  in  discipline  to  let  out  a 
poor  man  occasionally,  and  if  this  is  seldom  done  one  may  suspect  that  the 
foreman  is  not  strict  enough. 

Do  not  entrust  men  to  do  routine  work  without  supervision  and  inspec- 
tion, they  may  do  it  wrong  or  become  careless  if  they  realize  they  are  not 
watched. 

In  smelting  or  milling,  operations  are  carried  on  by  a  crew  who  work 
together,  each  being  responsible  for  his  assigned  duty.  All  are  directed 
by  the  foreman,  who  sees  that  operations  are  regular.  All  this  crew  are 
inside  men,  and  a  fresh  crew  replaces  the  preceding  one. 

PLANT-OPERATION 

Efficiency  of  Men. — In  smelting,  mechanical  appliances  are  increasingly 
in  use,  demanding  a  greater  investment  of  capital,  thus  reducing  the 
labor  per  ton  of  ore  treated.  This  makes  a  man's  work  less  strenuous,  and 
yet  better  paid.  Both  these  advantages  make  a  man  more  anxious  to 
retain  his  job,  and  so  he  is  more  dependable.  In  a  large  works  the  labor 
needed  per  ton  of  ore  is  less  than  in  a  small  one.  Steam-power  needs 
more  attendance  than  water-power,  and  is  on  the  whole  more  expensive. 


588  ORGANIZATION  AND  OPERATING 

In  starting  a  plant  into  operation,  a  list  of  places  and  occupations  of 
the  men  is  prepared,  so  that  men  who  are  chosen  may  be  quickly  assigned 
to  their  positions.  These  men  should  be  questioned  and  carefully  chosen. 
When  a  new  works  is  about  to  start,  skilled  men  often  apply,  and  they  may 
be  willing  to  do  common  labor  pending  the  starting  of  the  works,  and  in 
this  way  be  held  until  their  services  are  needed. 

Care  of  Men. — Provision  is  made  for  the  care  of  the  men  in  case  of 
sickness.  A  charge  of  $1  per  month  is  often  made  against  every  man,  and 
this  entitles  him  to  medical  attendance  and  care  at  the  hospital  in  case  of 
accident.  If  a  man  works  five  days  in  any  month,  the  $1  is  deducted  from 
his  pay  for  hospital  dues. 

MORALE  OF  INSIDE  MEN 

An  efficient  mill  or  smelter  man  is  proud  of  his  work,  and  to  encourage 
this  the  plant  should  be  kept  clean  and  orderly  on  all  shifts,  even  though 
this  adds  to  the  expenses.  He  should  have  training  in  repairing  his 
machines  and  not  have,  at  least  for  minor  repairs,  to  wait  for  the  repair 
gang. 

Where  those  in  charge  are  out  of  sympathy  with  their  men,  and  hold 
aloof  from  them,  it  conduces  to  a  larger  turnover,  that  is  to  men  leaving 
and  others  in  their  places  having  to  be  broken  into  the  work,  involving 
losses  in  so  doing.  Where  possible,  and  it  generally  is,  the  foreman  should 
notify  the  superintendent  of  his  intention  to  discharge  a  man,  thus  pre- 
venting its  depending  on  impulse  or  dislike. 

A  man  should  be  fitted  to  his  work,  and  if  he  fails  to  do  well,  he  may  be 
shifted  to  another  duty,  better  suited  to  his  capacity  or  tastes,  rather  than 
that  he  should  be  let  out.  Of  late,  for  a  large  works,  an  employment- 
specialist  tests  applicants  to  determine  not  only  their  fitness  but  also  the 
kind  of  job  at  which  they  can  do  their  best.  A  well-trained  man  should  be 
encouraged  to  stay.  The  practice  permitting  frequent  overtime  is  unsound, 
since  it  tends  to  excite  by  the  prospect  of  overtime  pay,  but  is,  in  the  long 
run,  exhausting. 

It  is  well  to  get  in  the  way  of  talking  over  works-methods  or  improve- 
ments with  the  men;  it  adds  to  their  efficiency  and  interest  in  their  work 
and  men  suggesting  improvements  should  be  rewarded  in  some  wafy.  The 
mill  man's  interest  is  increased  by  his  receiving  training  on  the  repair  gang, 
then  he  can  make  minor  repairs  himself. 

Contentment  is  increased  by  fair  wages,  bonuses,  comfortable  and 
attractive  surroundings,  yearly  vacation,  sick  pay,  and  medical  attention. 
Quarters  should  be  arranged  for  married  men,  and  for  single  ones  bunk 
houses,  where  they  can  sleep  undisturbed  when  off  shift.  If,  however, 
money  is  freely  spent  for  these  purposes  there  is  a  fear  that  the  men  may 


MODES  OF  PAYMENT  589 

think  the  company  is  rich,  so  why  not  strike  for  higher  wages.  There  is 
such  a  thing  as  being  too  conciliatory.  Even  where  the  company  can 
afford  it  they  should  rather  wait  for  the  men  to  ask  for  the  improvements, 
though  these  may  be  put  in  on  the  basis  of  increased  resultant  efficiency 
and  for  the  reason  that  it  is  the  custom  in  other  plants.  The  men  reason 
thus:  These  improvements  cost  much  money;  we  do  not  value  many  of 
them,  so  why  not  pay  their  cost  directly  to  us  as  wages. 

Sunday  closing  is  recognized  in  principle  as  giving  rest  and  a  change  in 
the  total  life  of  entire  plant  force.  To  interrupt  weekly  plant  operations 
and  to  bring  back  regularity  of  operation  on  Monday  may  involve  serious 
expenses  and  loss  in  extraction.  Moreover,  that  men  put  in  this  idle  time 
in  a  manner  detrimental  to  their  efficiency  is  another  argument  used.  Still, 
the  endeavor  in  any  plant  is  to  cut  out  operations  that  can  be  deferred  until 
Monday,  such  as  the  work  of  most  of  the  office  force. 

MODES  OF  PAYMENT 

Daily  wages  and  frequent  payments — especially-in  dealing  with  laboring 
men  of  the  primitive  races.  Monthly  payments  interfere  with  the  steady 
operation  of  the  works,  because  at  the  time  of  the  monthly  payment 
the  men  are  disposed  to  "  lay-off  "  to  spend  their  money.  To  overcome 
the  difficulty  two  methods  have  been  tried.  One  is  that  of  daily  pay- 
ments, by  which  a  man,  who  spends  his  money  when  he  gets  it,  has  only 
sufficient  to  supply  his  daily  wants  and  those  of  his  family,  and  none 
remaining  for  drunkenness  or  gambling.  The  other  system  is  to  pay  a 
wage  to  which  is  added  a  premium  that  increases  with  the  time  worked. 
It  is  paid  at  the  end  of  the  month  if  the  man  works  through  the  month,  but 
otherwise  not.  This  tends  to  keep  him  steadily  at  work. 

Also  careful  attention  must  be  paid  to  the  exactness  of  the  payroll, 
otherwise  dissatisfaction  may  result,  with  desertion  at  a  critical  time,  so 
that  the  works  suffer  from  labor  shortage.  This  difficulty  is  now  overcome 
by  the  use  of  indicator  clocks  by  which  the  time  of  entry  and  exit  is  punched 
on  the  employee's  card,  giving  an  unerring  record  that  he  can  reckon  up  for 
himself  at  the  end  of  the  month. 

Piece  Work. — This  is  the  payment  prevalent  in  steel  works  where  the 
men  are  paid  according  to  quantity.  The  great  difficulty  is  to  establish 
the  rate  which  shall  be  fair  for  both  sides.  If  the  employer  sees  that  the 
man  is  making  inordinate  profits,  due  to  his  speeding  up,  he  may  cut 
the  rate,  to  the  dissatisfaction  or  even  the  loss  of  the  man.  The  method 
becomes  complicated  where  several  men  work  together,  since  some  of  them 
are  more  efficient  than  the  rest.  A  subcontract,  however,  may  be  given 
one  man  who  hires  his  assistants.  In  practice,  in  certain  steel  works, 
wages  hare  increased  50  per  cent  to  60  per  cent  and  production  doubled. 


590  ORGANIZATION  AND  OPERATING 

Daily  Wages  and  Premium. — To  aid  in  this,  the  plan  of  progressive 
payments  has  been  instituted,  i.e.,  increased  payments  for  tons  in  excess 
of  normal  output.  At  the  Vieille  Montagne  zinc  works  several  systems  of 
payment  are  adopted,  but  the  mill  and  smelter  men  get,  in  addition  to  a 
fixed  wage,  a  premium  calculated  on  output,  and  another  premium  for 
unusual  energy.  In  some  cases  the  men  who  fire  the  furnace-blocks  get  a 
premium  based  on  the  time  their  furnace  lasts  without  repair.  In  other 
cases  thejretort  man  gets  a  premium  for  all  over  the  calculated  percent- 
age of  zinc  yielded.  Two-thirds  the  premium  is  paid  monthly,  the 
rest  is  retained  to  be  paid  at  the  end  of  the  year,  but  only  if  the  man 
has  "worked  regularly  throughout  the  year.  At  an  English  iron  works 
puddlers  are  paid  by  the  ton  with  a  premium  for  the  full  number  of  shifts 
during  the  week. 

Profit  Sharing. — The  system  of  participation  in  profits  on  the  part  of 
the  men  is  both  deceptive  and  dangerous.  It  is  possible  to  admit  officials, 
foremen,  and  specially  skilled  workmen  to  participation,  but  workmen  in 
general  are  not  fitted  for  the  change.  Everything  goes  well  as  long  as  the 
works  are  carried  along  at  a  profit,  but  in  bad  times  discontent  soon  breaks 
out.  The  system  is  not  favored  by  the  workmen  themselves.  They  are 
perfectly  willing  to  share  in  the  profits,  but  they  object  to  responsibility 
for  loss,  or  to  even  stand  for  the  creation  of  a  reserve  fund  to  cover  possible 
future  losses.  They  cannot  await  better  times,  nor  can  they  work  their 
turn  at  a  loss  in  order  to  retain  their  places.  Besides  this,  profits  appear 
too  remote,  and  they  cannot  understand  the  relation  that  exists  between 
the  work  and  the  annual  profit. 

Three  methods  of  profit-sharing  have  been  devised.  (1)  The  work- 
man gets  a  share  in  the  annual  cash  bonus;  (2)  This  bonus  is  kept 
back  for  a  specified  period  and  paid  him  together  with  the  accumulated 
interest;  (3)  At  the  Vieille  Montagne  works  the  one-third  portiofi  that 
comes  to  the  steady  worker  at  the  end  of  the  year  is  thus  distributed,  viz.,  a 
portion  of  it  is  at  once  given  him,  the  rest  invested  on  his  behalf.  The 
system  here  has  prevented  strikes. 

George  W.  Perkins,  a  well-known  financier,  proposes  as  a  correct  profit- 
sharing  plan  the  following: 

(1)  That  the  business  shall  first  of  all  earn  operating  expenses,  depre- 
ciation (a  serious  item  for  a  mine  plant  which  is  a  wasting  asset)  arM  a  fair 
return  of  4  per  cent  or  more  according  to  the  stability  of  the  venture  on 
the  capital  invested. 

(2)  That  all  profits  above  this  should  be  shared  on  a  percentage  basis. 
Thus  for  a  silver-lead  plant  the  labor  cost  is  given  at  25  per  cent. 

(3)  That  the  share  of  capital  should  be  carried  to  surplus,  the  share  of 
labor  should  be  distributed  to  them  as  a  security  (bond  or  stock),  of  the 
company.     The  employees'  share  of  the  profits  to  be  allotted  on  the  basis 


CAPITAL  REQUIREMENTS  591 

of  their  pay  and  that  each  employee  should  be  required  to  hold  his  security 
for  from  three  to  five  years. 

CAPITAL  REQUIREMENTS 

This  includes  not  only  capital  investment  and  the  funds  to  meet  future 
obligations,  but  also  the  working  or  quick  capital  called  the  liquid  assets, 
this  latter  being  that  needed  to  operate  the  plant.  It  includes  the  funds 
needed  for  purchase  of  ore,  flux,  fuels,  and  supplies,  to  pay  wages  and 
salaries,  and  to  meet  incidental  working  expenses.  This  includes  product 
bought  and  in  process  of  treatment.  Often  it  takes  thirty  to  sixty  days  to 
get  returns  on  finished  products.  It  often  happens  that  the  buying  com- 
pany or  a  bank  will  advance  money  to  the  seller  on  his  product  while 
awaiting  the  returns. 

Capital  Involved. — We  have  shown  that  a  custom  plant  needs  larger 
capital  to  carry  a  stock  of  ores  and  other  supplies,  but  with  a  works  plant 
the  ore  is  apt  to  be  but  a  short  time  to  process,  and  so,  especially  when  it 
ships  a  finished  product,  can  pay  its  way  with  little  trouble.  Capital  we 
understand  to  mean,  not  only  that  invested  in  plant,  but  that  needed  to 
carry  a  stock  of  ore,  of  supplies  as  well  as  to  pay  wages  as  they  come 
due.  Capital  when  invested  in  mines-plants  rightly  expects  a  large  return 
on  the  investment,  because  of  the  risk  involved,  that  the  property  may  not 
prove  up  as  expected.  For  we  must  remember  that  not  only  must  divi- 
dends be  large,  but  there  must  be  the  return  of  the  original  investment. 
When  a  mine  ceases  operation  its  plant  and  equipment  are  but  little  better 
than  scrap. 

It  has  been  the  fashion  to  complain  that  capital  is  in  the  hands  of  a 
few  greedy  rich  men.  In  truth,  however,  mining  stock  is  well  diffused 
throughout  the  community  and  much  of  it  has  little  value,  or  in  other 
words  the  investment  has  not  proved  to  be  profitable.  The  chance  of 
making  large  profits  has  enticed  many  into  such  investments  and  it 
has  been  said  with  truth  that  as  much  money  is  dropped  in  these  ventures 
as  has  been  taken  out  of  them. 

The  Costs  of  Production. — These  may  be  divided  into  prime  costs,  gen- 
eral costs  and  administration  costs.  Or  we  may  divide  them  into  working 
costs  and  overhead  costs.  Prime  costs  are  those  which  vary  according 
to  the  tonnage  put  through  and  which  cease  if  the  works  shut  down.  They 
are  also  called  flat,  actual,  or  direct  costs.  They  are  made  up  of  labor, 
motive  power,  fuel,  material,  supplies  and  repairs.  Under  material  is 
classed  what  is  used  over  and  over  again  in  specific  operations,  but  which  is 
gradually  consumed  as  zinc  in  the  Parkes  process  or  the  tank  acid  in  elec- 
trolytic copper  refining.  Supplies  include  tools  and  other  incidentals 
obtained  from  the  general  storeroom.  Repairs  include  not  only  the  labor, 
but  the  parts  needed  to  replace  those  worn  out. 


592  ORGANIZATION  AND  OPERATING 

General  Costs. — These  include  taking  care  of  the  items  of  interest, 
cartages,  lighting,  foremen,  watchmen,  miscellaneous  labor,  sampling,  the 
assay  and  chemical  laboratory,  testing,  etc. 

Administration  cost  takes  in  salaries,  office  expenses,  law  expenses, 
advertising,  traveling  expenses,  purchasing,  shipping,  selling,  taxes,  rents, 
etc. 

THE  ACCOUNTING  DEPARTMENT 

The  Council  of  the  Institution  of  Mining  and  Metallurgy  (British) 
unanimously  adopted  a  uniform  system  of  accounts,  the  outline  of  which 
will  be  followed  as  it  relates  to  metallurgical  works. 

These  accounts,  while  serving  for  guidance  in  the  distribution  of  expen- 
diture and  the  collection  of  revenue,  must  also  show  the  profit  or  loss  or 
the  financial  condition  of  the  undertaking  and  should  supply  the  manage- 
ment with  the  data  necessary  to  check  the  efficiency  of  the  administration 
and  to  point  out  how  useless  expenses  may  be  cut  down  or  revenue  increased. 

Capital  Expenditure. — Before  the  producing  stage  of  the  works  is 
reached,  all  expenditures  should  be  charged  to  capital  account  under  the 
following  heads:  (1)  Lands;  (2)  Machinery  and  Plant;  (3)  Buildings; 
(4)  Surface  works  as  reservoirs,  water  service,  railways,  sidings,  roads, 
power  line. 

After  the  plant  is  producing  all  other  expenditure  should  be  carried  to 
"  general  expenditure  account,"  to  be  distributed  proportionately  over  the 
remaining  heads  of  expenditure,  but  remembering  to  deduct  any  revenue 
that  may  have  been  received.  Also,  if  additional  property  is  purchased,  or 
additional  buildings,  machinery,  plant,  or  surface  works  put  up,  intended 
to  increase  production,  improve  recovery  or  decrease  costs,  then  these  items 
should  be  added  to  capital  expenditure  and  they  should  bear  the  share  of 
administration  or  general  charges.  If  any  machinery,  plant,  or  buildings 
should  be  entirely  superseded  or  replaced,  their  cost  should  be  taken  from 
capital  expenditure  and  be  charged  to  profit  and  loss.  In  case  the  item 
is  small  it  may  be  charged  in  full,  if  large,  then  by  installments  spread  over 
a  judicious  period. 

Temporary  or  Distribution  Accounts. — To  distribute  costs  as  evenly  as 
possible  it  is  well  to  open  temporary  accounts  in  order  to  spread  tlje  cost 
of  considerable  items  of  machinery  or  plant  replacement,  or  payments 
which  occur  annually  or  less  often  over  the  monthly  costs. 

Valuations. — Bullion,  concentrates  or  other  marketable  products  ready 
for  shipment  should  show  as  assets  on  the  balance  sheet  and  should  be 
credited  to  Revenue  Account  at  net  valuation  after  deducting  all  realiza- 
tion or  marketing  expenses.  Unfinished  products,  still  in  process,  should 
be  reckoned  at  cost,  provided  such  cost  is  less  than  market  value. 

Depreciation. — This  may  be  considered   as  it  affects  capital  expendi- 


ACCOUNTING  DEPARTMENT  593 

ture,  as  already  given.  Systematic  depreciation  is  theoretically  correct. 
It  is  the  amount  charged  annually  to  the  profit  and  loss  account,  accord- 
ing to  the  conservatively  estimated  life  of  each  item.  The  equipment  of  a 
mill  or  smelter  plant  may  depreciate,  for  example  at  the  rate  of  10  per  cent 
per  annum  upon  the  original  capital,  the  buildings  at  the  rate  of  5  per  cent. 
Thus,  even  if  kept  in  repair,  the  equipment  will,  at  the  end  of  ten  years, 
have  value  equivalent  to  scrap  and  may  be  practically  obsolete.  Depre- 
ciation may  be  divided  as  follows: 

Maintenance,  referring  to  the  wear  and  tear  on  equipment  and  buildings. 
It  varies  with  different  classes  of  equipment,  accidents  due  to  deteriora- 
tion, etc. 

Replacement. — Caused  by  wear  which  cannot  be  repaired  without 
replacement  of  the  worn-out  parts.  The  equipment  will  therefore  fall 
below  its  original  value. 

Obsolete. — Due  to  new  types  of  improved  equipment  which  are  neces- 
sary for  rapid  economical  production.  The  old  machine  can  only  be 
scrapped  and  has  no  value  except  as  old  iron. 

Neglect. — Even  if  properly  maintained,  the  equipment,  due  to  neglect, 
may  fall  below  its  actual  market  or  working  value. 

Inadequate. — Machines  may  become  too  small  to  be  of  service.  They 
often  render  co-operating  equipment  of  no  use,  since  they  are  inadequate 
to  serve  it.  Such  machines  have  a  value  if  sold  to  those  whose  operations 
need  them. 

Repair  and  Maintenance  Costs. — Suitable  provision  should  be  grad- 
ually made  out  of  income  for  new  buildings  and  equipment.  It  accumu- 
lates a  fund  for  new  plant  and,  at  the  same  time,  reduces  assets  to  some- 
thing like  their  true  value. 

Costs. — These  may  be  divided  into:  Flat  or  prime  costs,  otherwise 
specified  as  ore  treatment  or  reduction  charges,  and  second,  General  expense 
or  fixed  charges,  which  take  in  administration  and  general  charges,  realiza- 
tion charges,  taxes,  and  royalties. 

To  obtain  a  just  idea  of  costs,  where  mine  and  mill  are  hi  one,  we 
should  really  segregate  the  costs  of  mining,  concentrating,  and  reduction. 
The  costs  we  discuss,  however,  are  the  metallurgical  ones  and  it  is  expected 
that  they  should  be  used  in  connection  with  the  related  ones. 

Ore  Treatment  or  Reduction  Charges. — This  should  include  all  costs 
from  the  time  the  ore  is  delivered  at  the  plant  until  the  bullion,  metal, 
or  marketable  product  is  obtained,  and  may  be  divided  to  suit  circum- 
stances. In  it  should  come  (1)  power,  (2)  stores,  (3)  sampling,  (4)  assaying, 
(5)  maintenance  and  repairs,  (6)  salaries  and  payroll.  Maintenance  and 
repairs  should  be  segregated  either  on  the  basis  of  the  shop  expenses  or 
according  to  the  labor  employed  in  each  department. 


594  ORGANIZATION  AND  OPERATING 

ADMINISTRATION  AND  GENERAL  CHARGES 

These  include  (1)  consulting  engineer  and  general  manager's  fees, 
(2)  office  staff,  (3)  stationery,  postage  and  telegrams,  (4)  medical  and 
sanitary  expenses,  (5)  traveling  expenses,  (6)  fire  insurance,  (7)  em- 
ployers' liability  insurance,  (8)  hauling,  (9)  bank  charges,  (10)  auditors' 
fees,  (11)  legal  expense.  (Legal  expense  includes  questions  relating 
to  the  right-of-way  for  ditches,  flumes,  roads,  and  electric  power  line 
(see  page  549).  The  company  must  be  defended  against  those  farmers 
who  claim  alleged  damage  done  by  smoke.  Contracts  for  the  purchase 
of  ore  and  supplies  and  the  sale  of  products,  as  well  as  for  railroad  freight 
and  for  railroad  facilities  should  be  legally  drawn  up.  The  question  of 
royalties  and  labor  and  supply  contracts  are  settled  by  legal  advice. 
(Labor  disturbances  also  need  to  be  legally  handled.)  (12)  Directors'  fees, 
(14)  foreign  agency  expense,  (15)  interest  on  loans  and  bonds. 

Realization  Charges. — (1)  Hauling,  (2)  freight,  (3)  shipping  charges, 
agency,  and  commission,  (4)  sea  freight,  insurance,  etc.,  (5)  buying  and 
selling  expense,  (6)  advertising. 

Rents,  Taxes  and  Royalties. — These  include  (1)  rents  on  buildings, 
etc.,  (2)  tax  on  profits,  both  regular  and  excess,  (3)  public  taxes,  (4)  royal- 
ties. 

Reports. — The  annual  report  of  the  manager  should  show:  (1)  The 
quantity  of  ore  treated  in  tons  with  the  value  per  ton,  the  relative  values 
of  recovery  per  ton  in  each  department;  (2)  a  detailed  summary  of  working 
costs  subdivided  to  correspond  with  the  main  headings  and  subheading 
of  the  cost  sheets;  (3)  a  short  tabulated  statement  of  the  nature  of  and 
expenditure  upon  new  plant  and  equipment,  showing  sales  of  old  plant, 
if  any. 

Statements  as  to  the  quantities  and  values  of  supplies,  fuel,  fluxes,  ores, 
and  products  on  hand  should  be  certified  by  responsible  officials,  counter- 
signed by  the  manager.  The  taking  of  an  independent  inventory  is  advis- 
able from  time  to  time. 

Cost  Accounts  are  kept  for  the  purpose  of  determining  accurately  the 
cost  of  ore  treatment,  so  as  to  have  a  basis  for  treatment  charge;  to  judge 
how  well  operations  are  proceeding,  and  to  supply  data  for  plant  efficiency, 

TYPICAL  OPERATING  DEPARTMENT 

Silver-lead  or  Copper  Smelting  Works. — This  may  comprise  the  follow- 
ing responsible  men  under  charge  of  the  superintendent : 

Clerk  and  metallurgical  bookkeeper;  testing  engineers  in  control  of 
the  physical  and  chemical  laboratory  of  the  research  work  and  the  test- 
ing; engineers  and  mechanics  who  have  charge  of  the  power  plant  and  of 


OPERATING  DEPARTMENT  595 

machine  repairs;  foremen  to  attend  to  the  sampling,  to  the  yard  or  outside 
work,  to  the  roasters,  to  the  sintering  machines,  if  any;  and  to  the  blast- 
furnaces. 

Gold  Stamp  Mill  Producing  Concentrate. — The  force  may  include 
foremen,  mill  engineers  and  repair  force,  amalgamators,  feeders  and 
laborers. 

The  foreman  has  general  supervision  of  the  mill  and  looks  after  the 
handling,  cleaning,  and  retorting  of  all  amalgam  collected.  The  amalga- 
mators dress  the  chuck-blocks  and  plates,  and  keep  them  in  good  condition. 
They  set  tappets,  regulate  the  water-supply,  and  make  renewals.  The 
feeders  attend  to  the  uniform  feeding  of  the  batteries,  and  assist  the  amal- 
gamators in  renewals  and  at  the  clean-up.  A  good  feeder  is  a  valuable 
man  about  a  mill.  The  vannermen  attend  to  the  vanners,  or  concen- 
trating tables.  They  must  be  men  with  experience,  and  commonly  should 
first  serve  at  the  vanner  as  "  sulphide-pullers."  The  crusher-men  feed  the 
crushers  with  the  mine-ore  as  it  comes  to  the  mill.  Oilers  oil  the  machinery. 
Sulphide-puller?  remove  the  concentrate  or  sulphide  from  the  vanner  boxes, 
and  store  it  for  shipment.  Engineers  run  the  power-plant,  and  have  charge 
of  the  firemen.  Firemen  fire  the  boilers  and  remove  the  ashes.  Coal- 
passers  wheel  in  coal  from  the  coal-pile  to  the  boilers. 

On  repairs  there  are  carpenters,  with  laborers  to  help  them.  In  repairs 
on  the  vanners  there  is  a  special  vannerman  to  assist. 

Forty-stamp  Silver  Mill. — The  inside  labor  may  be  given  as  6  panmen, 
3  helpers,  and  15  tankmen  on  eight-hour  shifts. 

100-ton  Cyanide  Plant  Treating  Concentrates. — This  would  include 
the  superintendent  and  his  assistant;  the  accountant  and  the  chemist; 
the  foremen  for  each  shift  at  the  mill  and  for  each  shift  of  the  refinery; 
also  the  foremen  in  charge  of  repairs;  in  the  mill  the  solution,  filter  and 
pachuca  men;  in  the  refinery  the  refinery  men;  finally,  under  the  repair 
boss,  his  repairmen,  repair  helpers  and  the  common  labor  (often  called 
roustabouts). 

THE  PURCHASING   AND    SELLING   DEPARTMENT 

This  attends  to  the  purchase  and  delivery  to  the  works  of  ores,  fuel, 
fluxes,  and  general  supplies  for  repairs  and  renewals.  It  may  also  attend 
to  the  sale  of  the  products  of  the  works. 

The  Purchase  of  Fuel  and  Fluxes. — In  Utah  coke  may  be  quoted  at 
$12.60  per  ton,  f.o.b.  at  the  works,  at  Pittsburg  $5.  Attention  should  be 
paid  to  its  contents  in  moisture,  to  its  proportion  of  fines,  and  to  its  analysis, 
and  especially  to  the  amount  and  constitution  of  its  ash.  So  much  does  it 
vary  in  this  regard  that  the  buyer  should  be  well  informed  in  regard  to  the 
various  makes  of  coke  offered  for  purchase,  and  not  depend  on  the  price 
alone. 


596  ORGANIZATION  AND  OPERATING 

Western  coal  may  be  quoted  at  $5.50  per  ton  for  run-of-mine.  For 
certain  work  slack-coal  is  quite  suited.  The  favorable  qualities  are  low 
ash,  and  the  production  of  a  long  flame  in  the  reverberatory  furnace,  a 
quality  not  so  important  for  the  boiler. 

Limestone. — The  price  may  be  given  at  $2  per  ton  at  Utah  works.  For 
fluxing  it  should  be  low  in  silica.  It  should  not  be  friable  nor  contain 
much  fine. 

Iron  Ore. — Eight  dollars  per  ton  at  the  works  on  a  basis  of  47  per  cent 
iron  excess.  An  allowance  or  charge  of  20  cents  up  or  down  is  made  from 
this  figure. 

The  Purchase  of  Supplies. — These  consist  of  iron  and  steel,  castings, 
tools,  pipe  and  fittings,  oil  and  waste,  brick,  clay,  quicklime,  and  chemicals. 
The  list  price  of  most  of  these  are  given  in  the  catalogues  of  supply  houses 
and  discounts  from  the  list  are  given.  It  is  well  to  obtain  competitive  bids 
for  furnishing  these;  for  those  things  to  be  obtained  on  short  notice,  the 
buyer  can,  by  arrangement,  obtain  the  usual  discounts.  These  supplies 
are  kept  in  a  storeroom,  and  should  be  issued  by  the  supply  department 
only  on  a  written  order  from  the  foreman,  or  other  responsible  person  who 
needs  them.  In  this  way  it  is  known  where  they  are  to  be  distributed  on 
the  cost  sheets.  An  account  is  kept  of  all  supplies  received  and  issued,  so 
that,  from  it,  can  be  learned  how  much  and  when  to  order  such  material, 
to  maintain  the  stock.  It  is  detrimental  to  the  business  to  so  run  out  of 
supplies  as  to  cause  delays. 

Much  knowledge  is  required  in  the  purchase  of  supplies.  The  rule  is 
to  buy  when  prices  are  low,  or  on  a  rising  market,  but  only  in  small  quanti- 
ties on  a  falling  market,  and  to  obtain  the  best  discounts,  taking  care, 
however,  to  avoid  the  purchase  of  inferior  goods. 


CHAPTER  L 


PROFITS  AND  COSTS 


PROFITS 

Profits  from  the  operation  of  a  metallurgical  plant,  whether  a  mines 
plant  or  a  custom  plant  which  has  to  buy  its  ores,  may  be  defined  as  the 
difference  between  the  total  costs  and  the  returns  on  the  metal  product 
sales.  How  varied  are  these  costs  is  well  shown  in  their  enumeration  under 
head  of  "  Accounting,"  page  592.  Profits  may  be  increased  by  full  opera- 


$10.00 


0.00 


COSTS  AND  PROFITS 


WHEN   RUNNING  AT  THE  SPECIFIED 
PERCENTAGE  OF  CAPACITY 


$60,000 


50,000 


40,000 


30,000 


20,000 


10,000 


20*  40*  60*  80*  100* 

FIG.  330.— Costs  and  Profits. 

tion,  by  better  extraction  or  recovery  from  the  ore,  by  economy  of  treat- 
ment due  to  methods  permitting  a  saving  of  labor,  supplies,  or  fuel,  and 
by  faster  running  by  which  output  is  increased. 

The  above    graphic  table  shows   how  profits   increase    as  the    plant 
reaches  its  full  capacity;  it  also  shows  that,  in  this  particular  case,  profits 

597 


598  PROFITS  AND  COSTS 

cease  at  25  per  cent  of  capacity  and  again  that  the  profits  per  ton  increase 
as  full  capacity  is  approached. 

CUSTOM  SMELTERIES 

The  profits  of  a  custom  silver-lead  smelting  works  are  obtained  by  sub- 
tracting from  the  money  realized  by  the  sale  of  metals  recovered,  the  total 
costs  for  treatment,  freight,  refining,  interest  charges  and  selling  costs. 

Milling  Ores. — In  milling  the  calculation  remains  the  same  whether  the 
ore  is  highly  silicious  or  not. 

The  following  figures  represent  the  profits  of  a  company  owning  a  mine, 
the  Robinson  company,  on  the  Rand,  South  Africa: 

Gold  recovered  at  the  stamps $20 . 50 

Gold  recovered  by  cyaniding , 5 . 70 


Total  recovery $26.20 

Cost  of  mining $6 . 55 

Cost  of  milling 0.98 

Cost  of  cyaniding 0 . 97 

8.50 


Net  profits  per  ton $17 . 70 

Coeur  d'Alene  District. — Out  of  336,630  tons  of  ore  mined  at  the  Bunker 
Hill  &  Sullivan  mine  in  1906  there  were  shipped  to  the  smelting  works 
86,640  tons  of  concentrate  or  one  ton  in  3.84.  This  averaged  45.8  per  cent 
Pb  and  18.78  oz.  Ag  per  ton.  The  ore,  as  mined,  assayed  13.32  per  cent 
Pb  and  5.89  oz.  Ag  per  ton,  the  loss  by  concentration  being  estimated  at 
10.43  per  cent  Pb  and  17.06  per  cent  Ag,  or  11.96  per  cent  of  the  combined 
product.  Taking  the  average  prices  at  4.6  cents  per  pound  for  the  lead  and 
60  cents  per  ounce  for  the  silver,  we  find  the  costs  and  profits  per  ton 
as  below: 

Assay  value $15 . 78 

Mining,  milling  and  construction $2 . 43 

Freight,  treatment 3.71 

Smelter  deductions 3 . 08 

Mill  losses..  1.89           11.11 


Average  Profit $4.37 


INDEX 


Accessories  of  the  blast-furnace,  298 

-  (copper),  367 
Accounting  department,  592 
Accounts,  cost,  594 
— ,  distribution,  592 
Acid,  Bessemer  process,  321 
— lined  converter,  322,  323 

—  operation,  322 
— ,  open-hearth  process,  333 

—  parting  of  silver-gold  bullion,  278 

—  refractories,  31,  32 

—  treatment  of  zinc-box  precipitate,  183 
Action  of  machine  in  crushing,  50 
Administration  and  general  charges,  594 

—  department,  584 
Agglomeration  of  fine  ores,  286 
Agitation  of  slime,  157,  160 

—  treatment,  161 

Agitators,      combined      pneumatic      and 

mechanical,  160 
— ,  data  of,  187 
— ,  mechanical,  159 
— ,  pneumatic,  158 
Air,  combustion  in,  76 
Ajo  process,  432 

,  flow-sheet,  433 

,  leaching  tanks  for,  435 

Akins  classifier,  69,  70 
Alabama  pig  iron,  315 
Alaska  Treadwell  mill,  215,  216 
— ,  (tube-mill  circuit),  215 
Alkaline  earths,  action  of  in  slags,  482 
Allen  cone  classifier,  69,  70 
Alloy  steel,  347 
All  sliming  cyanidation,  157 
Alumina  in  slags,  482 
Aluminum  dust,  precipitation  by,  261 
Amalgam,  safe,  241 
— ,  treatment  of,  241     . 


Amalgamation  and  concentration  mill,  245 
-  of  silver  ores,  234,  244 

—  of  gold  ores,  123 

— ,  (plate),  123 

-  (patio  process),  249 
American  filtering  machine,  170 

-  ore  hearth,  464 
Ammonia  leaching,  438 
Anodes  and  cathodes,  451 

— ,  handling,  453 
— ,  sampling,  49 
Anthracites,  12 
Antimony  gold  ores,  145 

—  hi  silver-lead  smelting,  483 
Arbitration,  582 

Argall  roasting  furnace,  97 
Arsenic  in  silver-lead  smelting,  483 
Arsenical  gold  ores,  145 
Artificial  fuels,  11 
Association  of  employers,  583 
Augustin  process  of  silver  milling,  250 
Automatic  or  machine  sampling,  42 

B 

Bag  house,  488 
Ball  mill,  56 

,  Hardinge,  57 

in  closed  circuit,  56 

,  path  of  travel  in,  56 

— ,  proportions  and  efficiency,  56 
—  or  tube-mill  drive,  63 
Bar  screens  or  grizzly,  66 
Barrell  chlorination,  137 

— ,  plant  for,  141 
Base  bullion,  costs  of  refining,  511 

furnace,  501 

—  refining,  498 

sampling,  48 

,  punch,  48,  498 

— ,  selling  price  of,  511 


599 


600 


INDEX 


Base  bullion,  skimming  of,  503 
softening,  500 

—  metal  ores,  4 

Bases,  action  of  in  slags,  481 
Basic  Bessemer  process,  326 

—  copper  converter  lining,  402 

—  iron,  313 

—  open-hearth  charge  calculation,  335,  336 
,  chemistry  of,  337 

operation,  336 

process,  334 

—  refractories,  37 
Battery,  stamp,  125 
— ,  ten-stamp,  127 
Bedding  lead  ores,  467,  469 

—  ores,  40 

—  system,  Messiter,  361 
Beehive  coke,  319 

Belgium  zinc-smelting  furnace,  529 
Belmont  Milling  Co.  flow  sheet,  266 

mill,  265 

refinery,  268 

Belt  conveyor,  559 
Betts  process,  511 
Bessemer  pig,  315 

process  (acid),  321 

(basic),  326 

Black  copper  smelting,  359 
Blanket  or  canvas  tube,  114 
Blake  crusher,  51,  52 
Blast,  dry  air,  305 

—  or  pot  roasting  of  ores,  110,  286 

—  roasting  of  galena,  110 

-pot,  111 

Blast-furnace,  72,  287,  289,  290,  291,  292, 
298,  305 

—  accessories,  367 

—  burdening,  310 

—  combustion,  76,  305 

—  conditions  (copper),  369 

—  copper  matting,  369 

—  heat-balance,  309 

— ,  iron  and  accessories,  298 
— ,  — ,  operation  of,  300,  302 

—  irregularities,  302 

— ,  open  and  closed  top,  475 

—  operation,  371 

—  plant  for  copper  ores,  355,  356,  369 
for  iron,  287,  288 

— ,  showing  temperatures,  305 
— ,  silver-lead,  471,  472,  473 
— ,  — ,  products  of,  487 


Blast,  smelting  and  converting  plant,  406, 

407,  408 
— ,  —  of  copper  slag,  357 

— , sulphides,  360,  369 

— , precipitate,  188,  262 

—  tap  jacket,  474 

—  tuyere,  274 

—  vs.  reverberatory,  385 
Blende,  roasting  of,  519 
Blister  copper  refining,  442 
Blower,  positive  blast,  368,  369 
Blowing  engine,  299 

silver-lead  blast-furnace,  476 

the  iron  blast-furnace,  301 

Bone  ash,  37 
Bosh  and  molds,  8 
Boss  process  of  silver  milling,  243 
Bottoms,  treatment  of,  389 
Box,  zinc  or  extractor,  181 
Braun  disk  grinder,  42,  44 
Brick  kiln,  33 
-  making,  33,  35 

—  mold,  35 

—  repressing  machine,  36 
firicks,  32 

— ,  silica,  32 

Briquetting  flue  dust,  489 
^—  ores,  287 

Bromo-cyanide  process,  149 
Brown  agitating  tank,  158 

—  iron  ores,  285 

O'Hara  roasting  furnace,  97 

Brunton's  quartering  shovel,  42 

—  roasting  furnace,  107,  108 
Belt  elevators,  557 

Bins  and  pockets,  563 

Bitumite,  12 

Buchart  concentrating  table,  115 

Buggy,  charging,  551 

Burdening  the  iron  blast-furnace,  310 

Business  of  metallurgy,  578 

Butte-Duluth  mill,  431  £ 

—  process,  430 

Butter's  vacuum-leaf  filter,  168 
—  filter  frame,  169 
By-product  coke,  20,  21,  22 
plant,  22 


-{•^Calcination,  88 

Calculations  of  charge  in  pyritic  smelting, 
381 


INDEX 


601 


Calculations  of  charge   in    reverberatory 
smelting,  397 

silver-lead  smelting,  484 

of  mill  tonnage,  223 

Caldecott  diaphragm  cone,  69 

Callow  flotation  machine,  117 

Calorimeter,  Mahler  bomb,  80 

Calorimetry,  85 

Canvas  table,  114 

Capacity  of  roasting  furnace,  109 

Capital  expenditure,  592 

—  involved,  591 

—  requirements,  591 
Carbon  brick,  38 

—  in  pig  iron,  316 
Carbonate  iron  ores,  285 

Carriage  and  test,  English  cupelling  fur- 
nace, 509 
Cars,  hopper,  554 
— ,  industrial,  553 
— ,  side-dump,  553,  554 
— ,  transfer,  554 
Casting  copper,  8 
-  gold,  7 

—  iron  and  steel,  8 
-  lead,  8 

—  machine,  endless  mold,  414,  445 
—  Walker,  447 

—  silver,  7 

—  zinc,  9 
Castings,  steel,  346 
Cast  iron,  318 

—  mold,  318 
Centrifugal  belt,  507 
Chamotte,  35 
Characteristics  of  lead  ores,  461 

—  silver  ores,  231 
Chimneys  or  stacks,  78 
Charcoal,  16 

— ,  by-product,  17 

—  in  silver-lead  smelting,  484 
— ,  pig  iron,  315 

Charge    calculation    for    the    iron    blast- 
furnace, 310 

—  acid  open-hearth  process,  333 

—  basic  open-hearth  process,  339 
blast-furnace    matte  smelting,  375, 

381 

—  iron  blast-furnace,  312 

reverberatory  smelting,  397.  398 

— ,  reverberatory  copper  furnace,  393 

—  scoop,  zinc  smelting,  534 


Charges,  administration  and  general,  594 
— ,  realization,  594 
Charging  buggy,  551 

—  iron  blast-furnace,  294 

—  lead  ores,  567 

—  machine,  copper,  446 
— ,  open:  hearth,  333 

Chemical   reactions   of   the   basic    open- 
hearth  process^  337 

—  copper  converter,  404 

—  iron  blast-furnace,  479 

—  in  the  silver-lead  blast-furnace,  479 
Chemistry  of  roasting,  89,  92 

-  the  cyanide  process,  146,  264 
Chilian  mill,  60 
Chloridizing  blast  roasting,  249 

-  roaster,  MacDougall,  428 
— ,  Wedge,  424 

—  roasting,  88 

—  of  silver  ores,  246 

,  remarks  on,  248 

Chlorination  barrel,  137,  138,  139 

-  mill,  Goldfield,  136 

—  of  concentrates,  135 

—  copper  ores,  422 

gold  ores,  135 

— ,  ores  suited  to,  135 

—  or  Plattner  process,  133 

— .  precipitation  plant  for,  141 
Chromite  or  chrome  iron  ore,  34 
City  Deep  mill,  191 
Clarifying  solutions,  176 
Classification  and  definition  of  ore,  3 

—  of  iron  ores,  283 

—  metallurgical  operations,  5 

natural  solid  fuels,  13 

pig  iron,  314 

Classifier,  68 

—  Akins,  69,  70 

— ,  Allen  cone,  69,  70 
— ,  Caldecott  diaphragm  cone,  69 
— ,  Dorr,  56,  69,  71 
Classifying,  66,  68 

—  gold  mill  concentrates,  213 

—  of  gold  ores  for  milling,  122 
Clayey  gold  ore,  cyaniding,  189 
Cleaning  iron  furnace  gases,  295 
Clean-up,  129 

—  of  zinc  boxes,  182 

-  pan,  128,  129 

Coal-fired  reverberatory  furnace,  391 
Coals  for  roasting,  14 


602 


INDEX 


Coals  in  the  United  States,  composition 

of,  13 
Coarse  grinding,  55 

—  or  primary  crushing,  51,  265 
Cobalt  milling  practice,  274 
Coke,  18 

—  ash,  18 

— ,  costs  of  making,  23 

— ,  beehive,  19 

— ,  composition  of,  18 

—  in  silver-lead  smelting,  483 

—  oven,  by-product,  20,  22 

—  plant,  by-product,  22 

—  pusher  and  leveler,  22 
Colorado  lead  ores,  price  of,  483,  495 
Combined     mechanical    and     pneumatic 

agitators,  160 
Combustion,  75 

—  in  air,  76 

the  blast-furnace,  76,  77 

-  of  fuel,  81 
— ,  principles  of,  75 
— ,  temperature  of,  79,  82 
Comminution  of  ore,  62 
Comparative  agitator  data,  187 
Composite  costs,  575 
Composition  of  copper  matte,  372 
Compressed-air  locomotive,  552,  553 
Concentrate,  chlorination  of,  135 
— ,  sampling  of,  45 

—  treatment    at    Goldfield    Cons,    mill, 

218 
Concentrating  table,  Burchart,  115 

— ,  Wilfley,  115 
Concentration,  114,  265 

—  in  stamp  milling,  131 

—  prior  to  cyaniding,  143 
Concrete  floors,  unit  costs,  573 

—  foundation,  unit  costs,  572 
Condensers  for  zinc  smelting,  536 
Conical  mold,  8 

Coning  and  quartering,  41 
Consolidated  Langlaaghte  mill,  193 
Construction  of  plant,  550 
— ,  unit  costs  of,  572 
— ,  winter  work,  570 

Continuous  counter-current  decantation, 
166,  168 

—  grinding  pan,  237 

—  handling  of  materials,  557 
-  thickener,  Dorr,  163 
Conversion  of  white  metal,  405 


Converter,  acid-lined,  323 

—  and  mixer  building,  326,  341 
— ,  copper,  Smith-Pierce,  401,  403 
— ,  —  (upright  type),  401 

— ,  — ,  treater  for,  413 

— ,  electrolytically  operated,  323 

—  lining,  copper,  402 

-  (steel),  322 

Converting  copper  matte,  400,  403,  405 
-  leady  matte,  492 
Conveying,  365 
Conveyor,  belt,  559 
— ,  discharge,  559 
— ,  endless  chain,  561 
— ,  incline  to  level,  561 
— ,  screw,  559 

—  tripper,  561 

Copper  and  copper  ore  prices,  457 
bearing  gold  ores,  145 

—  blast-furnace  conditions,  369,  371 

-  plant,  355 

—  smelting  of  oxidized  ores,  355 

—  slags,  357 

—  casting,  9 

—  machine,  414,  445,  447 

—  charging  machine,  446 

—  converter,  400 

—  converting,  losses  in,  405 

—  electrolytic  tank,  450 

—  extraction  from  its  ores,  352 

—  furnace  slags,  374 

—  in  silver-lead  smelting,  483 

—  leaching,  A  jo  process,  432 
— ,  ammonia  process,  438 

— ,  Butte-Duluth  process,  430 

,  Henderson  process,  423 

— ,  Laist  process,  426 

-  plant,  440 

— ,  Rio  Tinto  process,  418 
— ,  Shannon  process,  421 
— ,  sulphuric  acid,  429 

—  matte,  373 

,  composition  of,  374 

-  converter,  400,  401,  402,  403 
— ,  converting,  400 

—  matting  blast-furnace,  363,  364,    369, 

370 

,  jackets  for,  365 

—  operation,  371,  373 
— ,  native,  352 

—  ores  and  their  treatment,  351,  417 
,  characteristics  of,  351 


INDEX 


603 


Copper  ores,  hydrometallurgy  of,  354 

— ,  pyrite  smelting,  354 

— ,  smelting  for  matte,  353 
— ,  properties  of,  353 

—  refining,  442,  449 

— ,  costs  of  operation,  455 
— ,  electrolytic,  449 

—  furnace,  443 

—  reverberatory  smelting,  387 
Costs  accounts,  594 

-  and  profits,  597 
— ,  composite,  575 
— ,  general,  592 
— ,  kinds  of,  593 

—  in  zinc  smelting,  538 

—  of  administration,  592 

—  agitation,  187 

—  concentration  in  milling,  132 

—  construction,  unit,  572 

—  copper  furnace  plant,  414,  415 

—  refinery,  455 

—  operation,  455 

—  cyaniding  on  the  Rand,  155 

—  silver-bearing  concentrate,  277 

—  dissolution  by  slime  agitation,  186 

—  filtration  or  decant ation,  187 

—  lead  ores,  494 

—  making  coke,  23 

—  metallurgical  plants,  571 

-  Homestake  plant,  156,  198 

—  operating  copper  works,  414,  416 

—  plant  and  equipment,  569 

—  production,  571 

—  of  pig  iron,  348 
-  steel,  348 

—  Rand  cyaniding  plants,  156 

—  rebuilding,  570 

—  refining  base-bullion,  511 

—  roasting,  109 

—  sampling,  45 

—  silver  ores,  280 

—  slime  plants,  186 

—  treatment,  Belmont-Tonopah  mill , 

227 

— ,  Homestake  mill,  227 
— ,  Modderfontein  mill,  227 

-  United  Eastern  mill,  211 
Cottrell  treater,  409,  413 
Counter-current  agitation,  165 
Cowper  hot-blast  stove,  297 
Crane  ladle,  406 

— ,  track  (and  grab  bucket),  555 


Crane,  traveling,  556 

Crowe  vacuum  process,  177 

Crusher,  Blake,  51,  53 

— ,  gyratdry,  52,  53 

— ,  Symons,  61,  62 

Crushing,  action  of  machines  in,  50 

—  and  screening,  dry,  64 

— ,  — ,  flow  sheet  for,  64 
— ,  coarse  or  primary,  51 
— ,  fine,  55 

— ,  grinding,    screening    and    classifying, 
50 

—  ores,  40 

— ,  principles  of,  50 
-  rolls,  58,  59 
Cupelling  furnace,  508 

—  rich  lead,  507 

—  silver  precipitate,  262 

Current    flow,    electromagnetic,    Walker 
multiple  system,  451 

— ,  testing,  454 
Custom  smelteries,  598 
Cupola  furnace,  73,  77 

— ,  combustion  hi,  77 
Cyanide  process,  133 

— ,  chemistry  of,  146 

—  solution,  strength  of,  146 
Cyaniding  concentrator,  221 
— ,  double  treatment  in,  113 

—  free  milling  ores,  189 

—  mixed  silver  ores,  268 
— ,  outline  of  process,  143 
— ,  sand  leaching  hi,  151 

— ,  silver-bearing  concentrates,  277 

—  silver  ores,  principles  of,  256,  264 
— ,  systems  of,  150,  223 
Cylinder  drier,  100 


Daily  wages  and  premium,  590 
Decantation  in  cyaniding,  157 
— ,  continuous  counter-current,  165 
— ,  costs  of,  187 
— ,  intermittent,  168 

Definition  of  metallurgical  thermochem- 
istry, 84 

—  straight  or  simple  ores,  3 
Definitions  and  classification  of  ores,  3 
Dehne  filter  press,  171 
Deister  concentrating  table,  115 
Department,  administration,  584 
—  operating,  585 


604 


INDEX 


Department,  purchasing  and  selling,  595 

Depreciation,  592 

Desilverizing  base-bullion,  276 

Desulphurizing  process  at  Nipissing,  276 

Dezincing  furnace,  505 

Ding's  magnetized  pulley,  566 

Dip  samples,  46 

Direct  process  of  reverberatory  smelting, 

390 

Disks  of  Symons  crusher,  62 
Disposal  of  pig  iron  at  iron  blast-furnace, 
304 

slag  at  copper  furnace,  384 

iron  blast-furnace,  303 

Distillation  of  zinc,  540 

—  of  zinc  ores,  528 
Distribution  accounts,  592 

—  of  wealth,  579 
Dolomite.  37,  482 
Dorr  agitator,  160 

—  bowl  classifier,  69,  71 

—  classifier,  56 

—  continuous  thickener,  163,  164 

,  showing  slime  zones,  164 

Double  strand  elevator,  558 

—  treatment  in  cyaniding,  158 
Dressing  the  zinc  boxes,  184 

—  the  plates  of  the  stamp  battery,  128 
Dry  air  blast,  305 

—  crushing  and  screening,  64 
flow  sheet,  65 

—  silver  milling  (Reese  River  process),  249 
Dryer,  cylinder,  100 

Drying  and  refining  the  gold  precipitate, 
184 

at  United  Eastern  plant, 

185 

silver  precipitate,  261 

Duties  of  foreman,  586 

—  superintendent,  585 

Duplex  and  electric  furnace  plant,    341, 
342 

—  process  of  steel  making,  340 
Dwight-Lloyd    sinter    roasting    machine, 

110,  112 

E 
Economic  situation,  579 

—  in  the  United  States  as  related  to 

metals,  580 

Economics  of  engineering,  579 
Edwards  roasting  furnace,  97,  100 


I    Efficiency  of  men,  587 
Electric  furnace  and  control  panel,  349 

duplexing  plant,  341 

building,  344 

for  steel  making,  342,  344 

smelting  of  silver  precipitate,  262 

showing  lining  and  bottom  connec- 
tion, 344 

—  lighting,  unit  costs  of,  573 

—  locomotive,  551,  552 

—  steel  making,  342 

—  transformer  and  substation  equipment, 

344 

—  trolley  removing  slag,  384 
Electrically  operated  converter,  323 
Electrolyte,  circulation  of,  454 

—  copper,  451 

Electrolytic  copper  refinery,  449 
refining,  449 

—  parting  of  silver  and  gold  bullion,  279 

—  refining  of  lead,  511 
-  tank,  450 

—  zinc,  542 
process,  542 

Electrostatic  Cottrell  treater,  409,  410 

—  recovery  of  smelter  dust,  410 
Elements,  heat  of  formation  of,  87 
Elevators,  566 

— ,  belt,  557 
— ,  feeding,  558 

— ,  single  and  double  strand,  558 
Elimination    of    impurities    (copper    con- 
verting), 405 

Employers,  association  of,  583 
Endless-chain  conveyor,  561,  562 

mold  casting  machine,  445 

Engineering,  economics  of,  579 
English  cupelling  furnace,  508 
Equipment,  obsolete,  593 

—  of  plants,  551 
Excavators,  555 
Expenditure  of  capital,  592 
Extraction  of  copper  as  chloricjp,  422 
by  leaching,  418 

—  smelting,  353 
—  from  ores,  352,  353,  418 
Extraction  of  gold  with  solvents,  133,  134 
silver  from  ores,  231 


Faber  du  Faur  retorting  furnace,  508 
tilting  furnace,  263,  508 


INDEX 


605 


Feed,  removal  of  wood  and  tramp  iron 

from,  565 
Feeders,  ore,  563 
— ,  — ,  and  shaking  screen,  564 
— ,  — ,  hammer,  565 
— ,  — ,  moving,  564 
— ,  — ,  reciprocating,  565 
— ,  — ,  rotary,  564 
— ,  — ,  traveling,  564 
Filter,  American  continuous  suction,  170, 

171 

—  frame,  Butters,  169 

— ,  Oliver,  continuous  suction,  170 
-  press,  Dehne,  171,  172,  173 

— ,  Kelly,  171,  172 

— ,  Sweetland,  173 

—  slime  treatment,  157 
Filters,  general  remarks  on,  175 
Filtration,  166 

— ,  costs  of,  187 

— ,  vacuum,  168 

Financial  crises  in  the  United  States,  583 

Fine  ores,  agglomeration  of,  286 

Finishing  the  sample,  45 

Fireclay,  firebrick  and  tile,  34 

Flame  temperature,  80 

Flotation,  116 

—  machine,  Callow,  116 

— ,  Janney,  116 

— ,  Minerals  Separation,  116 
Flowsheet  of  Alaska-Treadwell  mill,  216 

-  Hollinger  mill,  216 

-  Tom  Reed  mill,  208 

-  Waihi  Grand  Junction  mill,  271 
Flue  dust,  briquetting,  489 

— ,  treatment  of,  489 
Fluorspar  for  slags,  482 
Fore-hearth  or  settler,  portable,  366,  367 

— ,  stationary,  368 
Foreman,  duties  of,  586 
Foundry  pig,  315 
Fractional  selection,  41 
Free  milling  ores,  milling  of,  189 
Frenier  pump,  568 

—  installation,  568 
Frue  Vanner,  115 
Fuels,  11 

—  and  fluxes,  purchase  of,  595 
— ,  artificial,  11 

— ,  combustion,  81 

—  in  silver-lead  smelting,  483 
the  open-hearth  furnace.  332 


Fuels,  natural,  11 

— ,  oil  or  petroleum,  15 

Furnace  and  stoves  iron,  298 

— ,  blast,  72,  287,  298 

— ,  cupola,  73 

— ,  hand  reverberatory,  74,  96 

— ,  mechanically  operated  (roasting),  97 

— ,  natural  draft,  76 

— ,  reverberatory,  73,  74,  96 

— ,  revolving  roasting,  97 

-  shaft,  72 

— ,  softening,  501 

— ,  Wedge  roasting,  424 

— ,  wind,  72,  77 

—  zinc  smelting,  529,  530,  531,  532 


Galena  ores,  461 

Canister,  refractory,  32 

Gas  cleaning  at  the  iron  blast-furnace,  295 

— ,  natural,  15 

— ,  producer,  23 

— ,  —  plant,  25,  26,  27 

Gases,  metals  and  physical  constants,  6 

— ,  specific,  heat  of,  80 

Gasoline  locomotive,  552,  553 

Gate  for  ore  bin,  564 

General  arrangement  of  gold  stamp  mill, 

130 

—  an  iron  blast-furnace  plant,  312, 
313 

—  economic  situation,  579 

—  remarks  on  filters,  175 
Genesis  of  natural  fuels,  table  of,  11 
Gjiers  calcining  kiln,  286 

Gold,  amalgam  retorting,  129 

—  and  silver,  valuation  of,  122 

—  bars  or  ingots,  sampling,  48 

—  casting,  7 

—  mill  concentrates,  treatment  of,  213 
— ,  the  City  Deep,  191 

— ,  —  Victorious,  190 

—  milling  practice,  189 

—  ores,  amalgamation  of,  123 

—  and  classification  for  milling,  121 
— ,  antimony,  145 
— ,  arsenical,  145 
— ,  chlori  nation  of,  135 
— ,  copper-bearing,  145 
— ,  graphite,  145 
— ,  hydrometallurgy  of,  133 
— ,  occurrence  of,  121,  133 


INDEX 


Gold  ores,  prices,  227 

,  pyritic,  145 

,  silicious,  145 

,  smelting,  226 

,  talcose  or  clayey,  145 

,  tellurides,  145 

— ,  physical  properties  of,  121 

—  refining,  280 

—  tellurides,  121 

Golden  Cycle  flow  sheet,  201 

mill,  201 

Goldfield  chlorination  mill,  136 

—  Consolidated  mill,    concentrate   treat- 

ment at,  218 

Grab-bucket  and  track  crane,  555 
Grabs  and  excavators,  555 
Grading  ores,  5 

—  pig  iron,  315 
Graphite,  13,  15,  34 

bearing  gold  ores,  223 

Great  Falls  copper  converter,  401 
Grinding  and  classifying  on  the  Rand,  192 
— ,  coarse,  55 

—  pan,  continuous,  237 
Grizzly,  66 

Gyratory  crusher,  52,  53,  54 

H 

Hammer  feeder,  465 

Hand  reverberatory  furnace,  74 

—  sampling,  41 

Handling  of  materials,  continuous,  557 
Hardinge  conical  ball  mill,  57 
Heap  roasting,  92 

of  matte,  94 

Hearth  and  bosh,  iron  blast-furnace,  292 
Heat  balance  of  the  blast-furnace,  309 

—  evolved  in  roasting,  87 

Heats  of  formation  of  the  elements,  86 

,  table  of,  86,  87 

Hegeler  roasting  furnace,  97,  521,  523 

Hematite,  283 

Henderson  process,  423 

Hendryx  agitator,  160 

Heyl  &  Paterson  pig-casting  machine,  304 

High-grade  Nipissing  mill,  243 

Hoists,  industrial,  553,  554 

Holbeck  powdered  coal  system,  30 

Hollinger  mill,  205 

flow  sheet,  206 

Homestake  mill  (cyanide),  195 
sand  plant,  195,  196 


Homestake  mill  slime  plant,  197 
treatment  costs,  198 

—  plant  (cyanide),  cost  of,  156,  198 
Hopper  cars,  554 

Horizontal  copper  converter,  400 

,  Fierce-Smith  type,  403 

Hot  blast  stove,  295,  297 

—  metal  mixer,  321,  322 

—  solutions  in  cyaniding,  144 
Howard  mixer,  502 

—  press,  504 

Hughes  gas  producer  plant,  25 

Hy drome tallurgy  of  copper  ores,  354,  417 

gold  ores,  133 

silver  ores,  250 

Hyposulphite    lixiviating    of    silver    ores 
(Von  Patera  process),  254 


Impact  screen,  66,  67 
Impurities,  elimination  of  in  copper  con- 
verting, 409 

—  in  metals,  5 
Inadequate  equipment,  593 
Incline  conveyor,  561 
Indicating  pyrometer,  83 
Industrial  cars  and  hoists,  553 

—  locomotives,  551 

Influences  of  elements  on  pig  iron,  316 
Ingot  copper,  section  of  bar,  49 

—  mold,  324 

Inside  men,  morale  of,  588 
Intermediate  crushing,  55 
Intermittent  decantation,  168 

—  handling  of  materials,  551 
International  Smelting  Co.,  reverberatory 

smelting  and  converting  plant, 
413 

,  roaster  plant,  411,  412 

Iron  and  steel,  281 

casting,  8 

Iron  blast-furnace,  287,  289,  290,  291,  292, 
305  | 

and  accessories,  298 

,  burdening,  310 

charge  sheet,  312 

charging,  294 

— ,  chemical  reactions  of,  306,  307 

,  detailed  section,  291,  305 

plant,  287,  288,  289,  295 

,  showing  temperatures,  305 

with  automatic  charging,  290 


INDEX 


607 


Iron  ore,  action  of  in  slags,  481 

and  its  smelting,  283 

,  classification  and  occurrence,  283 

— ,  prices  of,  347 
—  roasting,  285 

Irregularities  of  blast-furnace  operation, 
302 


Jackson  centrifugal  pump,  567 
Janney  flotation  machine,  117 
Jones  sampler,  42 

K 

Kalgoorlie  district,  203 

Kelly  filter  press,  171 

Kennicott  plant  for  ammonia  leaching,  440 

Kiln  roasting,  86,  286 

Koppers  by-product  coke  oven,  22 


Labor  situation,  580 
— ,  union  and  non-union,  580 
— ,  unit  costs  for,  594 
Ladle  cars,  367 

—  crane  for  copper,  406 

— ,  worm-geared,  bottom-tapped,  324 
Laist  process,  426 

-  mill,  425,  427 
Lake  copper  refining,  444 
Large  copper  matting  blast-furnace,  369 
Leaching  (copper),  A  jo  process,  432 

— ,  ammonia  process,  438 

— ,  Butte-Duluth  process,  430 

—  ores,  417 

—  plant,  Henderson  process,  423 
— ,  Laist  process,  425,  426,  427 

—  the  sands  in  cyaniding,  153 

—  vats  in  cyaniding,  151 

Lead,  blast-furnace,  471,  472,  473,  475 
— ,  — ,  charge,  484 
— ,  casting,  9 

—  copper  matte,  491 
,  treatment  of,  491 

—  ores,  carbonates,  462 

— ,  classes  of,  461 

— ,  costs  of  smelting,  494 

— ,  oxidized,  462 

,  penalties  on,  496 

,  prices,  495 

,  properties  of,  461 


Lead  ores,  receiving,  sampling,  and  bed- 
ding, 467 

— ,  treatment  of  charge,  495 
— ,  properties  of,  461 

—  refinery,  499 
Liberty  Bell  mill,  199,  200 
Limestone  for  slags,  482 
Lining  of  copper  converter,  402 
Location  of  plants,  547 
Locomotives,  industrial,  551 
Long-hearth  reverberatory  roaster,  95,  96 
Loomis-Pettibone  gas  apparatus,  27 
Losses  in  copper  converting,  406 

zinc  smelting,  537 

M 

McArthur-Forrest  process,  133 
MacDougall  roasting   furnace,   101,   102 

104 

Machine  or  automatic  sampling,  42 
Machinery,  costs  of  creation,  573 

—  prices,  571 
Magnesite,  37 
Magnetite,  284 

Magnetized  pulley  for  removal  of  tramp 

iron,  566 

Mahler  bomb  calorimeter,  80 
Main  systems  of  filtering,  168 
Maintenance  and  repairs,  593 
Making  of  steel,  320 
Malleable  iron  pig,  315 
Manganese,  action  of  in  slags,  482 

—  in  pig  iron,  316 
Manufacture  of  steel,  320 

wrought  iron  by  puddling,  318 

Market  lead,  molding,  506 
Martin  sampling  machine,  47 
Materials  in  intermittent  handlkig,  551 
Matte  (copper),  373 
— ,  heap  roasting  of,  94 
— ,  leady  copper,  491 

— , converting,  492 

— ,  roasting,  108 

— ,  selling  price  of,  492 

—  smelting  charge,  375,  381 
(concentration),  383 

-  of  copper  ores,  354,  360,  371 
Mechanical  agitators,  159 

—  open-hearth  charging,  333 

—  roasting  furnaces,  97 

Melting  and  refining  Lake  copper,  444 
Men,  care  of,  588 


INDEX 


Men,  efficiency  of,  587 
Merrill  filter  press,  174 

— frame,  174 

installation,  173 

plate,  174 

—  precipitation  apparatus,  178 
process,  178 

Mercury  fed  to  the  stamp  battery,  126 
Merton  roasting  furnace,  97,  525,  526 
Messiter  bedding  system,  361 
Metallurgical  furnaces,  72 

—  operations,  classifications  of,  5 

—  thermo-chemistry,  84 

,  classification  of,  84 

,  units  of  measurement,  84 

Metals  and  gases,  physical  constants  for,  6 

— ,  impurities  in,  5 

— ,  molding  and  casting,  7 

Methods  of  treatment,  4 

Mexican  amalgamation  process,  249 

Mill  amalgamation  and  concentration,  245 

— ,  Belmont,  265 

— ,  Chilian,  60 

— ,  Consolidated  Langlaaghte  Co.,  193 

— ,  Golden  Eagle,  201 

— ,  Hollinger,  205 

—  iron,  314 

— ,  Liberty  BeU,  199,  200 

— ,  Nipissing  Co/s  low-grade,  274 

high-grade,  243 

— ,  Oroyo  Brownhill,  204  . 

—  samples,  46 

—  sites,  549 
— ,  stamp,  55 

— ,  Tom  Reed,  207 

— ,  tube,  64 

— ,  United  Eastern,  209 

— ,  Victor,  202 

— ,  Waiki  Grand  Junction,  270 

— , —  flow-sheet,  271 

— ,  Wasp  No.  2,  187 
Milling,  dry  silver,  249 

—  gold  ores  in  solutions,  133 

—  ores,  598 

—  practice  at  Cobalt,  274 
Mills,  typical  silver,  265 

Minerals    Separation    flotation    machine, 

116 
Mixed  gas  producer,  24 

—  ores,  definition  of,  3 
Mixer,  hot  metal,  321,  322,  341 
Mixing  pan,  324 


Modes  of  payment,  589 
Moisture  samples,  40 
Mold,  cast-iron,  8 
Molding  and  casting  metals,  7 
Multiple-hearth  furnace,  101 

N 
Native  copper,  351 

-  gold,  121 

Natural  draft  furnace,  76 
-fuels,  11 

— ,  genesis  of,  11 

—  gas,  15 

—  solid  fuels,  12 

— ,  classification  of,  12 
Neglect  of  equipment,  593 
Neutral  refractories,  31 
Newnam  ore  hearth,  465 
Nipissing  Co.'s  low-grade  mill,  274 
—  mill,  flow-sheet  of,  275        • 
wet  desulphurizing  process,  276 

—  mill,  high-grade,  243 
Nodulizing  iron  ores,  286 

O 

Obsolete  equipment,  593 
Occurrence  of  gold  ores,  115 

-  —  iron  ores,  283 
Oil  flotation,  116 
Oliver  continuous  filter,  170 
Open  and  closed-top  blast-furnace,  475 
Open-hearth  furnace,  327,  328,  329,  330 

-  building,  338,  339 
— ,  charging,  333 

— ,  —  machine,  333 

-  fuels,  332 

,  reversing  valves  for,  331 

— ,  steel-making  in,  326 

— ,  tilting,  332 

— ,  with  water-cooled  devices,  329 

-  process  (acid),  333 
— ,  (basic),  334 

,  (basic),  operation,  336        I 

— ,  calculation,  336 

,  chemistry  of,  337 

,  recarburization  in,  337 

Operating  department,  585 

,  cyanide  plant,  595 

.  forty-stamp  mill,  595 

— ,  gold  stamp  mill,  595 

— ,  silver  lead  or  copper  smelting,  594 
,  typical,  594 


INDEX 


609 


Operation  and  organization,  584 

—  of  basic  converter,  403 

—  copper  blast  furnace,  371 
-  plant,  587 

—  silver  lead  blast-furnace,  476 

—  the  acid-lined  converter,  322 
iron  blast-furnace,  300 

—  zinc  smelting  furnace,  434 
Ore  bins  and  pockets,  563 

-  feeders,  563 

—  gate,  564 

-  hearth,  464 

—  storage  and  supply,  563 

—  treatment,  593 
Ores  and  metals,  3 
— ,  base  metal,  4 
— ,  copper,  351 

— ,  definitions  of,  4 

— ,  gold,  occurrence  of,  121 

— ,  grading  of,  5 

— ,  mixed,  3 

— ,  preparation  of,  39 

— ,  receiving,  40 

— ,  storing,  46 

— ,  straight  or  simple,  3 

Organization  and  operating,  584 

Oroyo-Brownhill  mill,  204 

Outline  of  process  of  cyaniding,  143 

Overstrom  concentrating  table,  115 

Oxides  and  carbonates  of  copper,  352 

Oxidizing  roasting,  88 

Oxland  roasting  furnace,  97 


Pachuca  tank,  158 

Painting,  unit  costs  for,  574 

Pan,  clean-up,  129 

Parkes  process,  503 

Parting  silver  and  gold  bullion,  278,  279 

Path  of  travel  of  ore  particles,  56 

Patio  process  of  silver  amalgamation,  249 

Pattison  process,  510 

Payment,  modes  of,  589 

Penalties  on  lead  ores,  496 

Petroleum  or  fuel  oil,  15 

Phosphorus  in  pig  iron,  316 

Physical  properties  of  gold,  121 

Piece  work,  589 

Pig  casting  machine,  304 

Pig  iron,  314 

,  classification  of,  314 

,  costs  of  production,  348 


Pig  iron  disposal,  304 

—  grading,  315 

— ,  influence  of  elements  on,  316 

—  prices,  347 

—  sampling,  49 

— ,  smelting  for,  287 
Pittsburg  pig  iron,  315 
Plan  of  furnaces  and  stoves,  iron,  298 

-  United  Eastern  mill,  210 

Plant  for  duplex  and  electric  furnace,  341 

—  zinc  smelting,  533 

—  operation,  587 

Plants,  construction  of,  550 

— ,  equipment  and  their  costs,  546 

— ,  location,  547 

— ,  metallurgical,  costs  of,  571 

— ,  sites,  548,  549 

Plate  amalgamation  of  good  ores,  123 

— ,  dressing,  128 

Plattner  process,  133 

Plumbago,  34 

—  crucibles,  8 
Pneumatic  agitators,  158 
Portable  fore-hearth  or  settler,  366 
Positive  pressure  blower,  368,  369 
Pouring  ladle,  324 

—  slag,  373 

Power  needed  in  crushing,  50 

Practice  in  roll  crushing,  50 

Precipitate,  blast-furnace  smelting  of,  188 

— ,  gold,  184 

Precipitation  by  aluminum  dust,  261 

—  of  gold  from  solutions,  177 

silver  from  cyanide  solutions,  257 

Preliminary  construction  work,  570 
Premium  and  daily  wages,  590 
Preparation  of  ores,  39 
Pressure  filtration,  108,  171 
Prices  of  iron  ores,  347 

—  lead-silver  ores,  494,  495 

—  machinery,  571 

—  pig  iron,  347 

—  raw  materials,  575 

—  silver  and  silver  ores,  280 

-  steel,  347,  348 

—  zinc  ores,  539 
Principles  of  combustion,  75 

—  crushing,  50 

—  sampling,  39,  46 

—  the  hydrometallurgy  of  copper,  417 

-  silver,  250 

—  thermo-chemistry,  84 


610 


INDEX 


Principles    relating    to    the   refining    of 

metals,  5 
Producer  gas,  23 

,  costs,  28 

,  mixed,  24 

,  plant,  25 

,  simple,  23 

Production  and  costs,  591 

prices  of  lead-silver  ores,  494 

Products  of  silver-lead  blast  furnace,  487 
Profit-sharing,  590 
Profits  and  costs,  595 

—  of  treatment  of  silver-lead  ores,  496 
Properties  of  copper,  363 

lead  and  its  ores,  461 

Proportions  and  efficiency  of   ball   mill, 

56 
Puddling  furnace,  319 

—  process,  318 

,  reactions  in,  319 

Pug  mill,  490 

Pulp  sampler,  47,  48 

Pulverized  coal,  29 

distributing  system,  30 

-fired  reverberatory  furnace,  392 

in  silver-lead  smelting,  484 

Pumps,  Frenier  sand,  568 

— ,  high-pressure  triplex,  568 

— ,  three-throw,  567 

Purchase  of  fuel  and  fluxes,  595 

Purchasing  and  selling  department,  595 

Pyrite  gold  ores,  145 

—  smelting,  calculation  of  charge,  381 

of  copper  ores,  354,  377,  379,  383 

,  section  of  furnace,  379 

Pyrometer,  indicating,  83 

Q 

Quartering  an  ore  in  sampling,  41 

—  and  coning,  41 

—  shovel,  Brunton's,  42 


Rand  cyaniding  costs,  155 

—  plants,  costs  of,  156 

—  practice,  steel  tanks  for,  194 
Raw  material  prices,  575 
Raymond  roller-mill,  29 

Reactions  in  copper  reverberatory  smelt- 
ing, 397 

puddling  process,  319 

pyrite  matte  smelting,  378 


Reactions  in  sinter  roasting,  113 

Realization  charge,  594 

Rebuilding,  cost  of,  570 

Recarburization  in  open  hearths,  337 

Receiving  ores,  40 

— ,  sampling  and  bedding  lead  ores,  467 

Reciprocating  feed,  465 

Reclaiming    machine,    Robbins-Messiter, 

362 
Recovery  of  smelter  dusts,  410 

sulphuric  acid,  527 

Reduction  charges,  593 

Reese  River  dry  process  of  milling,  249 

Refinery,  lead,  499,  513 

— ,  Santa  Gertrudis,  260 

Refining  furnace  for  copper,  443 

—  in  a  cupelling  furnace,  184 

—  lead  and  base  bullion,  498 
—  of  copper,  442 

metals,  principles  relating  to,  5 

-  silver  precipitate,  259,  261 

—  with  potassium  bichromate,  184 

-  zinc,  540,  542 
Refractories,  31,  38 
— ,  acid,  31,  32 

— ,  basic,  31 

— ,  neutral,  31 

Refractory  materials,  properties  of,  131 

Regenerating  plant  (copper  refining),  453 

Regular  operation  of  the  iron  blast-furnace, 

302 

Rents,  taxes  and  royalties,  594 
Repair  and  maintenance  costs,  593 

—  force,  586 
Replacements,  592 
Reports,  594 

Repressing  machine,  brick,  36 
Requirements  of  capital,  591 
Research  work,  586 
Retort  and  furnace  for  silver  mill,  242 

—  manufacture,  536 

Retorting  furnace,  Faber  du  Faur^  508 

—  gold  amalgam,  129 
Retorts  for  zinc  smelting,  535     • 
Reverberatory  copper  smelting,  methods 

of,  385,  387,  390,  397 
,  Welsh  process,  387,  388,  389 

-  furnace,  73,  74,  327,  387 
Reverberatory  furnace,  open-hearth,  327 

—  roaster,  long  hearth,  95,  96 

—  smelting  and  converting  plant,  412,  413 
,  direct  process,  390 


INDEX 


611 


Reverberatory  smelting  furnace,  387 

— ,  charge,  calculations,  398 
-  — ,  coal-fired,  391 

— ,  method  of  charging,  392 

— ,  oil-fired,  394,  395 
,  pulverized-coal  fired,  392 

— ,  smelting  reactions,  397 

—  vs.  blast-furnace  smelting.  388 
Reversing    valves,    open-hearth    furnace, 

331 

Revolving  eccentric   gas-producer  plant, 
26,27 

—  screen  or  trommel,  68 
Rich  lead,  507 

,  cupelling,  507 

,  treatment  of,  507 

Ridge  roasting  furnace,  525,  526 
Rio  Tinto  process,  418 
Rules  of  works,  587 
Run-of-mine  ore,  size  of,  50 
Russell  process  for  silver  ores,  254 
Roaster  plant,  106 
Roasting,  88 
— ,  blast  or  pot-,  110 

-  blende,  519 
— ,  chemistry  of,  89 
— ,  chloridizing,  88 
— ,  coal  for,  18 
— ,  cost  of,  109 

—  furnace,  Argall,  97 

,  blende-,  521,  522 

,  Brown  horseshoe,  98 

,  Brown-O'Hara,  97 

,  Brunton,  107,  108 

,  cylinder,  98 

,  Edwards,  97,  100 

,  Hegeler,  97,  521,  523 

,  long-hearth  reverberatory,  95,  96 

— ,  MacDougall,  101,  104 

,  Merton,  97,  524,  525 

,  Oxland,  97 

,  Pierce  turret,  98 

— ,  Ropp,  97 

— ,  Wedge,  103,  105,  106 

— ,  Wethey,  97 

— ,  White-Howell,  97,  99 

—  furnaces,  capacity  of,  109 
— ,  heap,  92 

— ,  — ,  chemistry  of,  92 
— ,  heat  evolved  in,  87 
—  in  kilns,  88,  285 
the  Welsh  process,  389 


Roasting  iron  ores,  285 
— ,  losses  in,  108 

—  of  matte,  92 

—  ores  in  pulverized  condition,  94 
— ,  oxidizing,  88 

— ,  sinter,  88 

— ,  sulphatizing,  88 

— ,  triple,  113 

—  zinc  ores,  519 

,  chemistry  of,  519 

Robbins-Messiter     reclaiming      machine, 

362 

Roller  mill,  Raymond,  29 
Rolls,  crushing,  58,  59 
Roofing,  unit  costs  of,  574 
Ropp  roasting  furnace,  97 
Rotary  feeder,  464 

S 
Sample  finishing,  45 

—  grinding  mill,  42 

-  mill,  46 

— ,  moisture,  41 
Sampler,  Jones,  42 
— ,  pulp,  47 
— ,  tailings,  46 
— ,  Vezin,  43 

Sampling   concentrates,   tailings  and  ore 
pulps,  45 

—  copper  ingots  and  anodes,  49 
— ,  cost  of,  45 

— ,  hand,  41 

—  iron  ores,  46 

—  lead  ores,  467 

—  machine,  Martin,  47 

—  metals,  48 

-  mill,  43,  44 

-  ores,  40,  265 

—  containing  metallics,  45 

—  pig  iron,  49 

— ,  principles  of,  39,  46 

— ,  split  shovel,  41 

Sand  leaching  in  cyaniding,  151 

,  treatment  costs,  198 

—  plant,  Homestake,  147 

—  pump,  Frenier,  568 
— ,  refractory,  32 

San  Francisco  silver  mill,  266 
Santa  Gertrudis  refinery,  260 
Schedule  for  copper  and  copper-ore  prices, 

467 
Screen  impact,  66 


612 


INDEX 


Screen,  revolving,  or  trommel,  68 
— ,  sizes,  51 

—  tension,  67 
Screening,  66 

Screw  conveyor,  559,  560 
Scrubbers  for  stoves  and  boilers,  296 
Section  of  bar  or  ingot  copper,  49 
Selling  department,  595 
Semet-Solvay  by-product  coke  oven,  21 
Separation  of  slime  from  solutions;  166 
Settler,  eight-foot,  240 

—  or  fore-hearth,  portable,  366 

,  stationary,  368 

Shaft  furnace,  12 

Shafting,  pulleys,  and  belting,  unit  costs 

for,  574 

Shaking  screen  and  feeder,  564 
Shannon  Copper  Co.  process,  421 
Sharing  profits,  590 
Side-discharge  car,  554 

dump  car,  553 

Silica  brick,  32 

Silicates  of  copper,  352 

Silicious  gold  ores,  145 

Silicon  in  pig  iron,  316 

Silver  bars  or  ingots,  sampling,  48 

—  casting,  7 

— ,  distribution  of  in  a  bar  of  base  bullion, 
49 

—  extraction  from  ores,  232 

—  -gold  bullion  parting,  278,  279 
ores,  prices  of,  280 

—  -lead  bag-house,  488 

-  blast-furnace,  471,  472,  473,  477 
,  chemical  reactions  in,  479 

—  smelting,  467 
and  refinery,  513 

: plants,  location  of,  548 

,  pulverized  coal  for,  484 

,  slags  for,  480 

—  ores,  profits  of  treatment,  496 

-  works,  468,  471 

—  milling,  Boss  process,  242 

—  by  Augustin  process,  250 

—  hydrometallurgical  processes,  250 

—  Ziervogel  process,  251 
— ,  dry,  249 

— ,  wet,  235 

—  mills,  retort  and  furnace  for,  242 
— .  typical,  265 

-  with  tank  settling,  236 

—  ores  and  treatment,  231,  232 


Silver  ores,  amalgamation  and  concentra- 
tion, 244 

— ,  characteristics  of,  231 
,  chloridizing  roasting  of,  246 

— ,  cyanidation  of,  256,  268 

,  hyposulphite  process,  254 

,  Russell  process,  254 

—  precipitate,   drying  and  refining,  261, 

263 

,  melting  in  tilting  furnace,  262 

— ,  smelting  of,  262 
— ,  precipitation  of,  from  solutions,  257 

—  refining  furnace,  Faber  du  Faur,  263 

— ,  Monarch-Rockwell,  263 

— ,  Steel-Harvey,  263 
Simple  ores,  definition  of,  3 
Single-strand  elevator,  558 
Sinter  building,  470 

—  charge,  470 

-  roasting,  88,  110 
—  machine,  110 

,  reactions,  113 

Sintering  ores,  470 

Situation,  general  economic,  579 

Sizes  of  screens,  51 

Sizing  tests  on  sand  residue,  154 

Skimmer,  base  bullion,  505 

Skip  car,  555 

Slag  pot,  side-dumping,  478 

,  two-wheeled,  367 

Slags,  480 

— ,  actions  of  bases  in,  481 
— ,  ladle  and  locomotive  at  iron  furnace, 
303 

—  or  cinder  disposal  at  iron  furnace,  303 
— ,  pouring   (copper  blast-furnace  plant), 

373 
Slime  agitation,  157 

—  cake,  partly  removed  from  filter,  170 

—  plants,  costs,  Homestake,  198 
— ,  —  of,  186,  198 

— ,  refinery,  452  i 

— ,  separation  from  solution,  166 

—  treatment  by  filter,  157 
Sliming,  all,  157 
Smelteries,  custom,  598 
Smelting  by  the  Welsh  process,  388 

—  copper  ores,  353 

—  for  pig  iron,  287 

,  outline  of  process,  287 

—  on  the  ore  hearth,  463 

—  of  copper  slag,  357 


INDEX 


613 


Smelting  of  gold  ores,  226 
-  lead  ore,  463 

—  silver-lead  ores,  467 

— -plant  for  copper  ores,  355,  356 

—  silver  precipitate,  262 

—  to  black  copper  at  Union  Miniere  du 

Haut  Katanga,  357 

—  vs.  cyaniding  gold  ores,  226 

-  works  (silver-lead),  468,  471,  513 

-  zinc  ores,  519,  528 
Smith-Pierce  copper  converter,  401,  403 
Smoke  damage,  549 

Solution,  costs  of  by  slime  agitation,  186 

Solutions,  clarifying  of,  176 

— ,  precipitation  of  gold  from,  177 

Sorting  or  pickling  ore,  265 

Specific  heat  of  gases,  80 

Split-shovel  sampling,  41 

Stage  grinding,  50 

Stamp  battery,  125 

Stamp  mill,  55,  124 

— ,  using   amalgamation   and   concen- 
tration, 245 

—  with  plate  amalgamation,  124 
—  milling,  concentration  in,  131 
Stamping,  265 

Starting  sheets,  452 

—  the  copper  blast-furnace,  371 
Steam  locomotive,  551,  552 
Steel,  alloy,  347 

—  castings,  346 

— ,  costs  of  production,  348 

-  leaching  vat,  152,  192 

—  making,  320 

— ,  by  the  acid  Bessemer  process,  321 
— ,  —    —  duplex  process,  340 

—  in  the  open-hearth  furnace,  326 

-  prices,  347,  348 

-  rails,  345 

— ,  structural,  346 
— ,  tool,  346 

— ,  wrought  iron  and,  318,  345 
— ,  varieties  of,  345 
Storing  ores,  40 
Stove,  hot-blast,  295,  296,  298 
— ,  scrubbers  for,  296 
Straight  ores,  definition  of,  3 
Strength  of  cyanide  solutions,  147 
Structural  steel,  unit  costs  for,  574 
Sulphide  copper  ores,  blast-furnace  smelt- 
ing of,  360,  377 
Sulphides  of  copper,  351 


Sulphitizing  roasting,  88 
Sulphur  in  pig  iron,  317 
Sulphuric  acid  leaching,  420 

— ,  Ajo  process,  432 

— ,  Butte-Duluth  process,  430 
—  recovery,  527 
Superintendent,  duties  of,  585 
Supplies,  purchase  of,  596 
Sweetland  filter  press,  172 
Symons  disk  crusher,  61 
Systems  of  cyanidation,  150 


Tailing,  samplers,  46 

Talcose  gold  ores,  145 

Tank,  Brown,  58 

— ,  electrolytic,  450 

— ,  Pachuca,  58 

Tap-jacket,  blast  furnace,  474 

Taxes,  594 

Telluride  gold  ores,  145 

—  ores,  treatment  of,  199 

Telpher  mono-rail  transport,  555,  556 
Temperature  of  combustion,  79,  82 

-  flame,  80 

Temporary  or  distribution  accounts,  592 
Tension  screen,  67 
Ten-stamp  battery,  127 
Test  for  English  cupelling  furnace,  509 
Testing  or  research,  586 
Thermo-chemistry,  metallurgical,  84 
Thickening,  167 
Thimble  for  bag  house,  485 
Three-throw  plunger  pump,  567 
Tile,  34 

Tilting  furnace,  8,  414 
for  copper,  414 

— ,  melting  in,  262 

— ,  Monarch-Rockwell,  263 

— ,  open-hearth,  332 

— ,  Steel-Harvey,  263 
Tom  Reed  mill,  207 
Tonnage  in  mills,  calculations  for,  223 
Tool  steel,  345 
Tramp  iron,  removal  of,  465 
Transformer  and  substation  equipment, 

344 

Traveling  crane,  556 
Traveling  feeder,  464 
Traversing  bridge,  iron  smelting,  287 
Treatment  by  agitation,  161 

—  costs  of  lead  ores,  495 


614 


INDEX 


Treatment,  methods  of,  4 

—  of  amalgam,  241 

bottoms,  Welsh  process,  389 

copper  ores,  351 

-  flue  dust,  489 

gold-mill  concentrates,  213 

—  rich  lead,  507 

silver  ores,  232 

precipitate,  259 

tailings  for  acid  or  ammonia  leach- 
ing, 223 

telluride  ores,  199 

—  or  reduction  charges,  593 
Trent  agitator,  160 
Triple  roasting,  113 
Triplex  high-pressure  pump,  568 
Tripper  for  conveyor,  560,  561 
Trommel  or  revolving  screen,  68 
Tube  mill,  64 

circuit,  Alaska-Treadwell  mill,  215 

liners,  63 

—  milling,  265 

—  or  ball-mill  drive,  63 
Tuyere,  blast-furnace,  474 
Two-wheeled  slag  pot,  367 
Typical  gold  mills,  practice,  189 
-  silver  mills,  265 

U 

Underestimates  in  building,  570 
Union  and  non-union  labor,  580 

—  Miniere  du  Haut  Katanga,  smelting  at, 

359 
Unit  construction  costs,  572 

—  costs  for  concrete  foundations,  572 

erection  of  machinery,  573 

electric  lighting,  573 

excavation.  573 

floors,  573 

labor,  574 

masonry,  574 

painting,  574 

roofing,  574 

shafting,    pulleys,    and    belting, 

574 

structural  steel,  574 

ventilating  windows,   doors  and 

floors,  574 

United  Eastern  mill,  570 
United  Eastern  Mill,  costs  of,  211 

flow-sheet,  211 

plan,  210 


Units   of   measurement   in   thermo-chem- 

istry,  84 

Upright  copper  converter,  401 
Utah  lead  ores,  prices  of,  495 


Vacuum  filtration,  168 
—  -leaf  filter,  Butters,  168 

—  process,  177 
Valuation,  592 

—  of  gold  and  silver,  122 

Variations  of  costs  of  treatment  of  silver- 
lead  ores,  497 

Various  treatments  and  calculations,  223 

Vats  for  cyaniding,  description  of,  152, 
194 

— ,  steel  leaching,  154,  194 

— ,  — ,  for  double  treatment  of  cyaniding, 
154 

— ,  wooden  leaching,  152 

Ventilators,  windows  and  doors,  unit  costs 
of,  574 

Vezin  sampler,  43 

Victor  plant,  Portland  Gold  Mining  Co., 
202 

Victorious  gold  mill,  190 

Von  Patera  process,  254 

W 

Wage  scale,  575 

Waihi  Grand  Junction  mill,  270 

,  flow  sheet,  271 

Walker  casting  machine,  447 

—  multiple  system  of  current  flow,  451 
Wall-type  indicating  pyrometer,  83 
Wasp  No.  2  mill,  189 

Waste  wood  and  tramp  iron  removal,  465 
Water  jackets  for  copper    blast-furnace, 

365 

Wealth,  distribution  of,  579 
Wedge  blende  roasting  furnace,  522 

—  chloridizing  furnace,  424 

—  roasting  furnace,  522 

Welsh  process  of  reverberatory  smelting, 

387 

Wet  desulphurizing  process,  276 
Wethey  roasting  furnace,  97 
Wet  silver  mill,  236 

milling  with  tank  settling,  235 

White  briquetting  press,  490 
White-Howell  roasting  furnace,  97 

—  metal,  conversion  of  copper,  405 


INDEX 


615 


Wilfley  concentrating  table,  114,  115 

Wind  furnace,  72,  77 

Winter  work  in  construction,  570 

Wood,  14 

Working  costs  of  lead  ores,  495 

Workmen,  586 

Works,  rules  of,  587 

Worm  or  screw  conveyor,  560 


Ziervogel  process  of  extracting  silver,  251 
Zinc  boxes,  clean-up,  182 

— ,  —  and  treatment  of  precipitate,  258 

—  coating,  10 
-  distilling,  540 

— ,  grades  of,  540 
—  in  slags,  482 

—  ores,  517 


Zinc  ores,  reduction  of,  519 
-  — ,  roasting,  519 
— ,  — ,  chemistry  of,  519 

—  or  extractor  box,  181,  182 
—  prices  of,  539 

— ,  properties  of,  517 

—  refining,  540 

—  roasting  furnaces,  520 

—  smelting  furnace,  operation  of,  534 

—  furnaces,  529 

— ,  Belgium  type,  529 

,  regenerative  type,  532 

— ,  costs  of,  538 

,  losses  in,  537 

-  plant,  533 
,  prices  of,  539 

—  retorts,  534,  536 

Zones  of  the  blast-furnace,  300 


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